KR20050089867A - Recovering metals from sulfidic materials - Google Patents

Abstract

A process for recovering a precious metal from a sulfidic material comprises the steps of preparing an acidic aqueous halide solution having an oxidation potential sufficient to oxidise the sulfidic material and render the precious metal soluble in the solution, adding the material to the acidic aqueous halide solution so that the sulfidic material is oxidised and the precious metal is solubilised and separating the precious metal from the oxidised sulfidic material. In addition, a process for removing a contaminant from a contaminated sulfidic material comprises the steps of mixing the material in an aqueous solution wherein a multi-valent species of a relatively high oxidation state oxidises the contaminant to render it soluble in the solution, produces a contaminant refined material, and is reduced to a relatively lower oxidation state; and removing the contaminant from the solution whilst regenerating the multi-valent species to its relatively high oxidation state.

Description

How to recover metals from sulfur-containing soils {RECOVERING METALS FROM SULFIDIC MATERIALS}

The present invention relates to a method for recovering metals, in particular precious metals such as gold, from sulfidic materials. The process of the present invention can be applied to both contaminated and uncontaminated sulfur soils, which include relatively high carbon content (so-called "double refractory materials"), carbon free. Or sulfur-containing soils with a low carbon content (so-called "single refractory materials"). The term "relatively high carbon content" as used herein refers to the case where the carbon content in sulfur-containing soils is usually at least about 2% by weight.

A significant amount of sulfur-containing soils have been deposited throughout the world, including those for which economic recovery is feasible, especially precious metals such as gold and silver. For example, pyritic ores containing gold and / or silver and other precious metals such as platinum and platinum group metals are deposited and stored in large quantities.

Some of these deposits are contaminated with difficult-to-treat contaminants such as arsenic, antimony, bismuth or other heavy metals. In addition, ore treatment can be complicated when high concentrations of carbon are present, since carbon binds to precious metals such as gold and has a high affinity for such precious metals.

Conventional methods usable for oxidizing sulfur-containing soils include roasting, pressure oxidation (POx) and bio-oxidation (Biox). In POx and Biox methods, sulfate medium is usually used.

The method of roasting sulfur-containing soils presents an important problem because it emits sulfur-containing gases (so-called SOx gases) that are toxic to the environment. When arsenic is present in the ore, toxic substances such as arsenic trioxide are produced. For this reason, it is an international trend to avoid roasting sulfur-sulfur ores.

Pressurized oxidation processes of sulfur-containing soils are used to avoid the problems of roasting methods, but require high pressure (usually 30 bar or more) and relatively high temperatures (200 ° C. or more). In addition, pressurized oxidation processes are usually carried out in sulphate-based solutions.

U. S. Patent No. 6461577 discloses a production method for treating sulfur-containing soils containing arsenic, in which sulfur-containing soils are treated in a two-step Biox process to solubilize arsenic. The composition of the leaching process is complex because it uses bio-leaching bacteria. In addition, production methods are known to be slow.

U. S. Patent No. 4053305 discloses a leaching method for recovering copper and silver from sulfur-containing ores using a combination of ferrous chloride solution and pressurized oxygen. Copper dissolves during leaching, but silver is difficult to leach and remains a solid residue after leaching. Therefore, silver must be extracted from the residue using sodium cyanide, an environmentally hazardous leaching reagent.

U. S. Patent No. 4410496 discloses a leaching method for recovering copper, lead and zinc from a sulfur-containing ore by using a calcium chloride or barium chloride solution and pressurized oxygen in combination. Likewise, precious metals in the ore are not leached and must be extracted separately since they remain as solid residues after leaching.

U.S. Patent No. 4655829 discloses a leaching method for recovering metals from sulfur-containing ores including arsenic and antimony. In this method, bulky sulfur concentrates are prepared from arsenic-containing sulfur ore. The concentrate is slurried using excess calcium chloride solution. Once the concentrate has been produced, it is necessary to measure the total metal content and the concentrate composition. In order to prevent the formation of soluble arsenic compounds or toxic arsenic vapors, the concentrate is combined with an equilibrium solution slurry containing a predetermined concentration of copper, lead, zinc or mixtures thereof in the form of sulfides of the metal. At this time, the concentrate and the equilibration solution slurry is combined to obtain a reaction slurry having a metal content determined such that the molar concentration of arsenic and antimony in the mixture is approximately equal to the molar concentration of copper, lead and zinc in the range of about 60-40 or 40-60. Form. Once the mixture has been properly equilibrated, it is heated and vented under pressure to oxidize the metal into soluble components. In other words, equilibrium is essential so that no soluble arsenic compounds or toxic arsenic vapors are produced.

If there is a simple hydrometallurgical method, it would be advantageous to recover precious metals, especially gold, from sulfur-containing soils.

1 is a schematic representation of the POx and Biox process according to the prior art, compared to the preferred method of the present invention (IRGP) for recovering precious metals from sulfur-containing soils.

FIG. 2 is a schematic view of a process diagram for a first method for recovering precious metals (gold) from contaminated sulfur-containing soils (non-ferrous-FeAsS).

3 and 4 are graphs showing the relationship between the amount of gold and iron extraction and the Eh of the solution at various stages of IRGP.

FIG. 5 is a schematic view of a process diagram for a second method of removing contaminants from sulfur-containing soils while simultaneously recovering precious metals from sulfur-containing soils.

FIG. 6 schematically shows a process diagram for a preferred method for removing contaminants from a single refractory sulfur-containing soil and recovering precious metals from the sulfur-containing soil.

FIG. 7 is a schematic illustration of a process diagram for a preferred method for removing contaminants from a double refractory sulfur containing soil and recovering precious metals from the sulfur containing soil.

FIG. 8 is a graph depicting the relationship of various first stage (arsenic) leaching solution parameters to time (duration of reaction) for the methods of FIGS. 6 and 7.

FIG. 9 is a graph showing the relationship of the various second stage (pyrite) leaching solution parameters to time (reaction duration) for the methods of FIGS. 6 and 7.

Summary of the Invention

The first invention of the present application provides a method for recovering a noble metal from sulfur-containing soil, comprising the following steps:

Preparing an aqueous acid halide solution having an oxidation potential sufficient to oxidize the sulfur-containing soil and make the precious metal soluble in the solution;

Adding the sulfur-containing soil to the aqueous acid halide solution to oxidize the sulfur-containing soil and solubilize the precious metal; And

Separating said precious metal from oxidized sulfur-containing soils.

The inventors unexpectedly found that sulfur-containing soils can be oxidized in one step simultaneously with the solubilization of the precious metal if sufficient oxidation potential is maintained in the acid halide solution.

In addition, the inventors have found that when sulfur-containing soils are contaminated with arsenic, antimony, etc., arsenic and the like can be leached and precipitated at the same time, solubilizing the precious metal in one step without using a pre or initial solution equilibrium step.

Thus, in a second aspect the present invention provides a method for recovering precious metals from contaminated sulfur containing soil comprising the following steps:

Preparing an aqueous acid halide aqueous solution having a oxidation potential sufficient to oxidize the sulfur-containing soil and solubilize the precious metal in the solution and to precipitate arsenic;

Adding the sulfur-containing soil to the aqueous acidic halide solution to oxidize the sulfur-containing soil, solubilize the precious metal and precipitate the arsenic; And

Separating said precious metal from oxidized sulfur-containing soil and precipitated arsenic.

The method of the first invention and the second invention is distinguished from POx and Biox in that halide is used instead of sulfate-based leaching solution.

The inventors noted that halides (similar to cyanide) form strong complexes with precious metals such as gold, thereby facilitating dissolution of the precious metal and subsequent recovery of the precious metal by carbon adsorption. However, because halides are weaker ligands than cyanide, the inventors have developed a method to achieve a cyanide-like noble metal solubility potential by sufficiently high oxidation potential (Eh) in an acidic environment (preferably pH <3).

The process of the present invention is advantageous in that performing in a closed or recirculated manner yields economical benefits (eg, simplicity, low energy consumption, preservation of mass balance, etc.). We also recover precious metals from any sulfur-containing soils, including ores and concentrates that are difficult to process with other methods, such as double refractory soils (eg, carbon-containing arsenite) with relatively high carbon content. It turns out that it can be applied.

After the solution containing the noble metal is separated from the oxidized sulfur-containing soil and precipitated arsenic (if present) through a solid-liquid separation step, the precious metal from the solution is recovered in a metal recovery step, preferably using one or more carbon-containing columns. It is preferable to recover by activated carbon adsorption. Preferably, after adsorbing the noble metal on activated carbon, the carbon is removed and calcined to recover the noble metal, or the carbon is eluted with a cyanide solution and the eluted product is subjected to an electrolytic step, followed by molten salt electrowinning. Recover. At this time, the method of the present invention is clearly distinguished from the conventional method. In the conventional method, it is necessary to cyanide the oxidation reaction residue in order to extract the precious metal (gold), so that a separate delicate leaching circuit is required. In the present invention, the cyanide leaching process is unnecessary because the precious metal is already solubilized at the time of leaching. In addition, many environmental agencies currently require additional decomposition of cyanide, particularly in environmentally sensitive areas.

In the case of uncontaminated sulfur soils (eg, single refractory pyrite uncontaminated with arsenic, etc.), the oxidation of sulfur-containing soils is generally carried out in one step. In the case of contaminated sulfur-containing soils (e.g., single or double refractory pyrite contaminated with arsenic and / or carbon, etc.), the oxidation reaction of sulfur-containing soils is generally carried out in two stages, but solubilization of precious metals takes place in the first stage.

Typically, the solution is recycled to the sulfur-containing soil oxidation step after the metal recovery step. The metal recycling step is preferably carried out after the solid-liquid separation step in the in-line form and before the step of recycling the solution to the sulfur-containing soil oxidation step. The term "inline" means a step provided as part of a solution circuit (ie, a "circuit" formed by recycling the solution to a sulfur-containing soil oxidation step). Moreover, metal recovery methods other than carbon adsorption methods, for example, ion exchange, a solvent extraction method, etc. can also be used.

In the case of double refractory ores containing carbon, it is necessary to have additional metal recovery steps (i.e., separate from the solution recycle circuit) in order to recover precious metals passing with the solid material obtained from the sulfur-containing soil oxidation step. . This extra step is necessary because some precious metals (eg gold) are not solubilized directly through the oxidation step with carbon. A separate metal recovery step uses conventional roasting or metal dissolution methods, and in some cases precious metals remaining in the solid material roasted using a leaching method (eg, a solution obtained from a sulfur-containing soil oxidation step) after the roasting method. (E.g., gold).

Typically, the precious metal to be recovered is gold, but may be silver, platinum or other platinum group metals and are metals which are economically justified by recovering them.

The aqueous halide solution is most preferably a soluble metal halide solution, usually a solution having a halide concentration of about 8 mol / liter. The halide is preferably chloride, but bromide or halide mixtures such as chloride and bromide can also be used.

The method is carried out in such a way that the metal in the dissolved metal halide can act as a polyvalent species. At this time, as the polyvalent species, both those having a relatively high oxidation state and generally having a relatively low oxidation state reduced during the oxidation reaction are selected so as to precipitate upon oxidation of the sulfur-containing soil. The polyvalent species can be regenerated in a relatively high oxidation state, after which the recycled polyvalent species is recycled to the sulfur-containing soil oxidation step to advantageously participate in subsequent oxidation reactions. Economical benefits (e.g., mass equilibrium) may be achieved by regenerating the multivalent species during the leaching step (s) to recycle the recycled species to the sulfur-containing soil oxidation step, preferably as part of a closed or recycle process. Conservation, simplicity, low energy consumption, etc.).

Usually, the metal in the metal halide solution is copper, but iron or the like may be used. Any one of these polyvalent species effectively acts as an electron transfer agent. For example, in a solution recycled to the sulfur-containing soil oxidation step, the metal is present in a relatively high oxidation state (eg Cu (II) or Fe (III)) and after its oxidation reaction in its relatively low oxidation state. (E.g. Cu (I) or Fe (II)). In the leaching stage, the polyvalent species are usually present in pairs (ie, high and low oxidation states). However, other polyvalent species, optionally cobalt, manganese, vanadium or the like may also be used.

When the sulfur-containing soil is arsenopyrite, by adjusting the oxidation potential, arsenic can be leached into the solution in the first leaching step. However, once leached, it is desirable to adjust the pH of the solution so that arsenic precipitates in the form of scorodite. In addition, in the case where the sulfur-containing soil is arsenite, it is preferable to leach the pyrite component in the second step, and in this step, it is preferable to maintain the arsenic in the ferric arsenate precipitation state by adjusting the pH of the solution. Thus, arsenic is discharged from the process along with the solid residue in the solid-liquid separation step and does not prevent the recovery of precious metals.

In the case of uncontaminated single refractory pyrite soils, the sulfur-containing soil oxidation step usually comprises a single leaching step, in which the precious metal is solubilized as the sulfur-containing soil is oxidized.

Each leaching step can be carried out co-currently or countercurrently, in which one or more vessels can be used.

Preferably, the entire solution obtained from the first leaching step is fed to the second leaching step.

If sulfur-containing soils are contaminated with, for example, arsenic, the first leaching stage generally requires that the sulfur-containing soils are under Eh sufficient to leach contaminants and solubilize precious metals (e.g. gold), preferably about 0.7-0.8 volts ( Contact with the solution). In the Eh of this solution, the pyrite component of the sulfur-containing soil is substantially not leached. Preferably, the pH of the solution in the first leaching step is 1 or less but about 0.5 or more so that the contaminants may be precipitated immediately after elution. In the first leaching step, the temperature of the solution is preferably about 80-105 ° C, more generally 80-95 ° C.

In the case of uncontaminated sulfur soils (in this case using a single leaching stage), or in the second leaching stage used to leach the pyrite component of contaminated sulfur soils, the sulfur-containing soil is usually Eh sufficient to leach pyrite. , Preferably with a solution having an Eh of about 0.8-0.9 volts. Similarly, the pH of the solution must be less than or equal to about 0.2 or greater so that the contaminants can precipitate immediately after leaching. In addition, in the case of leaching pyrite, the temperature of the solution is higher than or equal to the temperature in leaching arsenite, and is usually about 95-105 ° C.

To obtain a higher solution of Eh in a single or second leaching step, it is necessary to add additional oxidants such as oxygen, air, chlorine gas, hydrogen peroxide and the like. To achieve an optimal solution pH to maintain contaminants in precipitated form and regenerate secondary copper ions, an acid such as sulfuric acid and / or a base such as sodium carbonate may be added to a single or second leaching step to raise the pH. There is a need. Otherwise, arsenic and iron will be solubilized without precipitation. At this time, in the single leaching step or the second leaching step, the pyrite component of the sulfur-containing soil may produce sufficient or excess sulfuric acid. As another method, hydrochloric acid or another acid which does not cause chemical interference with the method may be used.

On the other hand, the separated solution after the leaching step is fed to the noble metal recovery step, and the separated residual solid is usually discarded.

Preferably, after the precious metal recovery step, the concentration of the chemical species is adjusted by removing (precipitating) ferric sulfate using a solution adjustment step. Usually, in this step limestone and calcium carbonate are added to the solution to form a hematite / gypsum precipitate, which is then filtered and treated with the solid residue from the leaching step (s). However, the step of removing the ferrous component is controlled in a manner that controls the amount of limestone addition to maintain some iron in the solution, thus preventing the precipitation of copper in the form of cupric (ie, copper at low pH). This is because iron precipitates and buffers the pH at the time of precipitation, acting as an inhibitor against copper precipitation).

Although it is preferable to filter the solid residue from the solution in the solid-liquid separation step, other methods may be used, such as solid / liquid settling, solution evaporation, centrifugation methods and the like.

If high concentrations of carbon are present in sulfur-containing soils (e.g. 2-20% by weight of carbon), surfactants such as blinding agents are added during the sulfur-containing soil oxidation step to noble metals (e.g. gold) It is advantageous to prevent adsorption on carbon in the soil. The blinding agent is generally at least one organic solvent, including kerosene, phenol ethers and the like. Alternatively, activated carbon may be added to the solution to adsorb gold in advance. Through the use of blinding agents or activated charcoal, the need for a separate metal recovery step to separate precious metals that may be passed with carbon as a solid residue can be eliminated.

The most advantageous use of the present invention relates to a method for recovering precious metals from pyrite and concentrates, which naturally occur in pyrite soils in the mined state, where contaminants are usually arsenic, antimony, bismuth, mercury, cadmium and the like.

Other economically important metals such as copper, nickel, zinc, lead and the like can also be recovered further in the process of the invention. In addition, for certain applications, it may be desirable or necessary to recover the contaminants themselves. For example, if a contaminant is economically valuable or environmentally harmful, it may be necessary to recover it from contaminant deposits (eg, if the contaminant is antimony, bismuth, cadmium, etc.).

The method of the second invention of the present application is used when sulfur-containing soil is contaminated with arsenic, antimony and the like. In this method, the contaminants are leached and precipitated, while simultaneously solubilizing the precious metal, without the need to use a pre or initial solution leaching step. For example, in some applications where the pollutant must be recovered separately (eg where the pollutant is economically valuable), or as an alternative to the second invention herein, the step of contaminant precipitation may be carried out. It may be advantageous to separate from the leaching step.

Thus, according to a third invention of the present application, there is provided a method for removing contaminants from contaminated sulfur soils comprising the following steps:

The sulfur species are mixed in an aqueous solution, oxidizing the contaminants with a polyvalent species having a relatively high oxidation state solubilizing the contaminants in the solution, thereby producing the sulfur-containing soils from which the contaminants have been removed. Reducing to a low oxidation state; And

Recycling the polyvalent species to its relatively high oxidation state while simultaneously removing contaminants from the solution.

This method also recovers metals, particularly precious metals such as gold combined with contaminated sulfur-containing soils. In addition, by regenerating the polyvalent species at the same time as removing contaminants, the process of the present invention can advantageously be carried out in a closed or recirculating manner, thus providing economic advantages such as simplicity, low energy consumption, preservation of mass balance, and the like. Entails.

Other uses, such as when contaminants need to be removed prior to the roasting or smelting process of conventional sulfur-containing ores, or as an alternative to the second invention herein, contaminate the contaminant precipitation step. It may be desirable to separate from the material leaching step.

Accordingly, the fourth invention of the present application provides a method for removing contaminants from contaminated sulfur containing soil comprising the following steps:

Mixing sulfur-containing soils in an aqueous solution with an oxidation potential controlled to substantially oxidize only the pollutants and solubilize them in solution to produce contaminant-free soils; And

Separating the solution from the contaminated desulfurized soil.

By controlling the oxidation potential, the method of the fourth invention herein advantageously facilitates the removal of contaminants in the future as it can advantageously keep the contaminants in soluble form (for example a separate precipitation step).

For example, in the case where the sulfur-containing soil is arsenite and the pollutant is arsenic, the oxidation potential can be adjusted to oxidize and solubilize arsenic in the first leaching step and not to oxidize pyrite. In addition, in the method of the third invention and the fourth invention, once the arsenic is solubilized and separated, the remaining pyrite component can be more strongly oxidized in a subsequent leaching step (eg, a second leaching step).

As used herein, the term "contaminated sulfur soils" includes soils from which contaminants have been completely removed, but meets environmental standards to the extent that they can be processed (eg in roasting and smelting units) or at disposal. This includes soils with pollutant concentrations low enough to be sufficient. The methods of the third and fourth invention herein are typically used to treat pyrite or concentrate, including arsenic, antimony, bismuth, mercury and cadmium as contaminants. These pollutants are naturally occurring in the pyrite soils as they are mined. The method of the third and fourth inventions of the present application is also applicable to treating ores and concentrates that are difficult to treat, such as ubiquitous or especially double refractory ores with high carbon content.

In the third and fourth inventions herein, the contaminants are removed from the solution by precipitation in a separate precipitation step by introducing an oxidant into the solution. It may be advantageous for the oxidant to oxidize the polyvalent species to a relatively high oxidation state at the same time. The solution may then be recycled to the leaching stage after the contaminants are precipitated and removed and the multivalent species is regenerated to a higher oxidation state.

In the precipitation step, the pH of the solution is maintained at about pH 1.5-3. The pH of the solution is usually maintained by adjusting the amount of oxidant and / or alkaline reagent supplied to the solution. When it is necessary to add an alkali reagent, alkali salts such as calcium carbonate, calcium oxide, sodium carbonate, sodium bicarbonate and the like are usually added.

In the precipitation step, the oxidant generally oxidizes the polyvalent species at the same time as it precipitates by oxidizing the contaminants in a relatively poorly soluble form (eg, by oxidizing arsenic from the +3 oxidation state to the +5 oxidation state). The oxidant may be air, oxygen, chlorine gas, hydrogen peroxide gas, or the like. In pyrite ore, pollutants usually precipitate as iron / pollutant oxide forms (eg iron arsenate if the pollutant is arsenic).

After precipitation of the contaminants, the solution returns the Eh and pH to the levels required for contaminant leaching so that the solution can be recycled to the leaching step. This can be achieved, for example, by adjusting the amount of oxidant addition after the pollutant precipitation.

In the process of the third and fourth invention herein, the contaminants can be oxidized and eluted into solution through a one-step or multi-step leaching process. Generally, the leaching method comprises a first leaching step of adjusting the oxidation potential to substantially oxidize only the contaminants to make the contaminants soluble in the solution, and an oxidation potential to oxidize the sulfides in the sulfur-containing soil from which the contaminants have been removed. And a second leaching step of increasing. At this time, most of the contaminants may be oxidized and solubilized in the first leaching step, and the remaining contaminants may be oxidized in the second leaching step.

In addition, each leaching step can be carried out co-currently or countercurrently, in which one or more vessels can be used.

Generally, contaminant-containing sulfur-containing soils are separated from the solution after the first leaching stage and fed to the second leaching stage. In addition, the solution is usually separated from the contaminated sulfur-containing soil after each leaching step, so that the contaminants can be removed from the solution by precipitation in the usual precipitation step.

In the process of the third and fourth inventions herein, when the sulfur-containing soil is pyrite (eg ubiquitous or other contaminated pyrite), in the first leaching step the contaminant is generally acidic with a pH of 1 or less. In aqueous solution, the contaminants are oxidized at a temperature of about 105 ° C. or less, under Eh of a solution sufficient to oxidize the solution into the solution but substantially without eluting pyrite, typically under Eh of about 0.7-0.8 bulter (based on SHE). In the second leaching step, the pyrite soil is usually oxidized in an acidic aqueous solution having a pH of 1 or less, under an Eh of a higher solution sufficient to elute the pyrite, usually at an temperature of about 105 ° C. or less, under an Eh of about 0.8-0.9 volts. Let's do it. In order to obtain a higher solution Eh in the second leaching step, an oxidant such as oxygen, air, chlorine gas, hydrogen peroxide gas, or the like may be added to the solution. In addition, an acid such as sulfuric acid may be added if necessary.

In the second leaching step, sulfuric acid, hydrochloric acid or other acids without chemical interference to the process need to be added to maintain the pH of the low solution to oxidize pyrite and to solubilize the arsenic, which usually remains in the +5 oxidation state. There may be. However, it may not be necessary to add acid (for example, if sulfur present in the ore or concentrate is oxidized to produce sufficient sulfuric acid in the solution).

As with the methods of the first and second inventions herein, the solution recycled through the leaching and precipitation steps is usually a dissolved metal chloride solution with a chloride concentration of about 8 moles / liter, in which the metal is a multivalent chemical Act as species (as defined above with respect to the first and second inventions herein).

As with the methods of the first and second inventions herein, when high concentrations of carbon are present in the pyrite (e.g., 2-20% by weight of carbon), surfactants such as blinding agents may be In addition to the solution, it may be advantageous to prevent dissolved metals (especially precious metals such as gold) from adsorbing onto the carbon in the soil. The use of blinding agents may rule out the need for a roasting step to separate the precious metal from carbon.

Accordingly, the fifth invention of the present application provides a method for recovering precious metals from sulfur-containing soils by treating contaminated sulfur-containing soils having a relatively high carbon content:

Leaching the sulfur-containing soil in an aqueous solution to leach the metal into the solution, during which the carbon in the sulfur-containing soil is shielded to prevent the adsorption of precious metals on the carbon; And

Recovering said precious metal from solution.

The term “relatively high carbon content” refers to the concentration of carbon present in sulfur-containing soil, usually on the order of about 2-20% by weight.

Carbon may be masked with a blinding agent as described above. Other features of the method of the fifth invention herein are as described in the first to fourth inventions.

After the contaminants are precipitated out and the polyvalent species is regenerated to a relatively high oxidation state, the solution is generally recycled to the leaching stage. Since polyvalent species have been regenerated in their original (prior to leaching) oxidation state, they are easy to oxidize and leach later.

In the methods of the third and fourth inventions herein, a metal recovery step can be performed to recover metal present in the soil from which contaminants leached and / or remaining with the contaminants into the solution.

For example, in the case of double refractory ores containing carbon, it is possible to recover metals (e.g. precious metals such as gold) present in the remaining soil that has been decontaminated, such as adsorbed on carbon after the final leaching stage. A metal recovery step may be necessary to do this. Also, in the case of double refractory ores, the metal recovery step may consist of a conventional roasting step or a smelting step, since contaminants have been substantially removed from the sulfur-containing soil during the leaching step. If desired, the chlorine or cyanide leaching method can be used after the roasting step to recover the metal remaining in the roasted solid material (for example if the metal is a precious metal such as gold).

Alternatively, or in addition, the solution in the leaching step before the contaminant precipitation step (i.e. immediately before the contaminant oxidation step and the precipitation step) or after the contaminant precipitation step (i.e. immediately after recycling to the contaminant precipitation and oxidation step) An in-line metal recovery step may be necessary to remove the metal leached into it. The term "inline" refers to a step located on a solution recycle circuit. The inline metal recovery step usually involves the adsorption of the metal in solution onto the carbon in the carbon column, usually on activated carbon. Alternatively, other metal recovery methods such as ion exchange and solvent extraction can be used.

Typical metals recovered in the process of the third and fourth inventions herein include precious metals such as gold, silver, platinum or other platinum group metals, which are economically justified metals. However, other economically important metals may also be recovered, alternatively or additionally, such as copper, nickel, zinc, lead, and the like. In addition, in certain applications of the third and fourth invention herein, it may be desirable or necessary to recover the contaminants themselves. For example, if a contaminant is economically valuable or environmentally harmful, it is desirable to recover it from the contaminant precipitate (eg, for contaminants such as antimony, bismuth, cadmium, etc.). If the contaminant constitutes the "metal" to be recovered, the contaminant recovery step may be performed additionally or in general after the contaminant precipitation step.

Prior to recovering the metal in the method of the third and fourth invention herein, a number of soil separation steps are typically performed to separate the soil from which the contaminants have been removed. At this time, the solution is generally concentrated after the first leaching step by concentrating and separating the soil from which the contaminants have been removed. After the second leaching step, the contaminant-removed soil is filtered out of the solution, but other separation methods such as solid / liquid settling, solution evaporation, centrifugation and the like may also be used.

Thus, after each of the first leaching step and the second leaching step, the separated solution is passed through to the contaminant recovery step, while the separated refined soil is fed to the metal recovery step (e.g. in the case of double refractory pyrite). It needs to be discarded.

In addition, in the method of the third invention and the fourth invention of the present invention, after the pollutant precipitation step, the pollutant separation step is performed to remove the contaminants from the solution and then the solution is leached (or before the in-line metal recovery step). Recycle. At this time, the solid / liquid separation step is used after the contaminant precipitation step, and it is easy to use a filtration method or another separation method.

Hereinafter, the present invention will be described with reference to preferred embodiments based on the accompanying drawings, but the embodiments shown in the drawings are merely exemplary, and are not intended to limit the protection scope of the present invention.

Before describing the preferred method of the present invention in detail with reference to the examples, the preferred method of the present invention will be described in general by comparing with the prior art POx and Biox method with reference to FIG.

A preferred method according to the invention is hereinafter referred to as Intec Refractory Gold Process (IRGP). This method was developed as an alternative to using halides to recover gold from refractory sulfur-containing soil deposits. Known treatment methods for such deposits generally float the pulverized ore to produce a concentrate, which is then treated to oxidize the sulfur-containing soil mainly to sulphate, and finally to use cyanide from the oxidation residue. To extract gold.

As a conventional method which can be selected in order to oxidize a sulfur containing mineral, the roasting method, the pressurized oxidation method (POx), and the production method (Biox) are mentioned. It is schematically shown in FIG. 1 by comparing IRGP with wet smelting procedures (POx and Biox) currently in use. IRGP and wet smelting POx and Biox are distinguished in that they use halides, not sulfate media, in IRGP. While gold is insoluble in sulfates, cyanide-like halides form strong complexes with gold to facilitate gold dissolution and subsequent recovery of gold by adsorption on activated carbon. Since halides are weaker ligands than cyanide, an acidic environment (pH <2) and higher solution temperature and potential (Eh) are used to achieve the same gold extraction efficiency.

In order to treat refractory sulfur-containing soils, using a halide medium with controlled solution oxidation potential, it is possible to oxidize arsenic and sulfides and dissolve gold. After separating the gold-containing solution from the oxidized sulfur-containing mineral slurry, the dissolved gold can be recovered by adsorption on activated carbon, which is subsequently calcined or eluted with cyanide to finally dissolve the gold by melt electrolysis. Recover as. Unlike conventional methods, IRGP does not need to cyanide the oxidation residue to extract gold, which cyanide reaction decomposes a separate dedicated leaching circuit and in some cases residual cyanide. It requires an extra cost to make it.

There are a number of factors that can provide gold-containing ore refractory, as shown in Table 1 below.

type Causes of fire resistance characteristics Glass Physical confinement to silicates, sulfides, carbon, etc. Occlusion Passivation by the formation of chemical layers chemistry Formation of gold-containing compounds such as gold telluride and aurostibnite substitution Elemental substitution by gold in the mineral lattice, eg "Soluble" gold in pyrite absorption Adsorption of Molten Gold by Activated Carbon Materials in Ore Pulp

IRGP was specifically developed to treat concentrates produced from refractory ores belonging to the last two categories of Table 1 above, namely "substituted" and "adsorbed" categories. Most of the gold present in the world falls into these two categories and is named ubiquitous and pyrite depending on the presence of iron sulfide, which occurs separately or more often in mixed form. IRGP may also be applied where "activated carbon" is present in the ore.

Hereinafter, the IRGP method and the chemical principle will be described below for a method of treating refractory gold concentrates containing minerals of the following types.

1. Ubitic iron

2. ubiquitous + pyrite

3. ubiquitous + pyrite + carbon

Ubiquitous  Chemical description of the oxidation reaction

If arsenic is present in refractory gold concentrates, it is mainly in the form of arsenic (FeAsS). Gold is often referred to as a solid solution rather than natural gold because it is "locked" in arsenite as a lattice bonded species. Therefore, in order to release the gold, it was necessary to completely destroy the ubiquitous lattice.

The breakdown of the arsenite lattice in IRGP is achieved by chemical oxidation reaction according to Scheme 1 below.

FeAsS + 2O ₂ → FeAsO ₄ + S

Because of the very low solubility of oxygen in the treatment liquid, oxygen did not directly oxidize the arsenite, but it worked through several intermediate steps.

Oxygen is supplied directly from air sparged under atmospheric pressure upon leaching, and is first used to produce soluble oxidants in the form of first copper ions (Cu ²⁺ ) according to Scheme 2 below.

2Cu ⁺ + ½O ₂ + 2H ⁺ → 2Cu ²⁺ + H ₂ O

The reaction occurred at the interface between the bubble and the treatment liquid. Subsequently, the first copper ions oxidized the arsenic iron in accordance with the following Scheme 3.

FeAsS + 7Cu ²⁺ + 4H ₂ O → H ₃ AsO ₄ + Fe ²⁺ + S + 5H ⁺ + 7Cu ⁺

The reaction product of ferrous and first copper was subsequently further oxidized by air sparged according to Scheme 2 and Scheme 4 below.

Cu ²⁺ + Fe ²⁺ → Cu ⁺ + Fe ³⁺

In the presence of ferric iron ions, the fly ash readily formed insoluble iron arsenate according to Scheme 5 below.

H ₃ AsO ₄ + Fe ³⁺ → FeAsO ₄ + 3H ⁺

Iron arsenate formed an electrolyte with a high chloride content under the operating conditions used in the IRGP method, which was usually crystalline and stable in the surrounding environment and could be easily separated.

Since there was always a small amount of background iron concentration in the treatment liquid, the action of the Cu ²⁺ / Cu ⁺ phase was supplemented by the Fe ³⁺ / Fe ²⁺ pair. The potential attainable under the influence of Cu ²⁺ and Fe ³⁺ was 850 mV (SHE basis) in the presence of oxygen. This potential was sufficient to dissolve gold because gold was stabilized by forming chloride complexes according to Scheme 6 below.

^{3Cu 2+ + Au + 4Cl - →} AuCl - 4 + 3Cu ⁺

If bromide is present in the treatment solution (eg, intentionally added), a gold-bromide complex was also formed according to the following scheme:

^{3Cu 2+ + Au + 4Br - →} AuBr - 4 + 3Cu +

Oxidation was carried out at a temperature of 90-95 ℃ in 8M chloride electrolyte containing the Fe ^{+ 3} ions in the 20-40 g / l of Cu ^{2 +} ions and 2-5 g / l.

Chemical description of pyrite oxidation

Oxidation reaction of pyrite (FeS ₂ ) in IRGP was accomplished through the same series of intermediate reactions used in the arsenite oxidation reaction, according to the overall scheme as shown in Scheme 7 below.

4FeS ₂ + 15O ₂ + 2H ₂ O → 8SO ₄ ^2- + 4Fe ³⁺ + 4H ⁺

It has been found that the sulfur of pyrite is oxidized to sulphate in some way, whereas the sulfur of arsenite is oxidized only to the elemental state.

Since pyrite is more fire resistant than arsenite, finer particle sizes were used to obtain acceptable reaction rates as described below. However, each pyrite sample showed variable reactivity, which is believed to be affected by the substitution of arsenic for some of the sulfur in the crystal lattice. Such pyrite is often referred to as arsenic-containing pyrite, the higher the arsenic contamination, the closer the reactivity of pyrite is to the reactivity of the actual arsenite with an As / S ratio of 1.

To according to Scheme 8, the reaction was carried out in the same treating solution as used in the same manner as arsenopyrite at a temperature of 90-95 ℃ through the Cu ^{2 +} / Cu ⁺ couple, arsenopyrite oxidation.

FeS ₂ + 7Cu ²⁺ + 4H ₂ O → SO ₄ ^2- + Fe ²⁺ + 8H ⁺ + 7Cu ⁺

Cu ⁺ and Fe ^{+ 2} have been further oxidized by oxygen to be sprayed according to the scheme 2, and 4. The ferric sulfate formed was precipitated into hematite and gypsum by adding limestone under pH about 1-1.5 according to Scheme 9 below.

4SO ₄ ^2- + 2Fe ³⁺ + 2H ⁺ + 4CaCO ₃ → Fe ₂ O ₃ + 4CaSO ₄ + 4CO ₂ + H ₂ O

The amount of limestone added was adjusted to maintain the soluble iron concentration at 2-5 g / l to prevent precipitation and loss of copper residues in the second copper form by leaching.

Concentrate crushing size

Fruits for use in IRGP were usually obtained in sizes ranging from 70-100 microns or less in the 80% size range. As a result of the test, when the concentrate was regrind to a finer size (according to the characteristics of each concentrate), the reaction rate was significantly increased, and the first example described later used a regrinding method. If ubiquitous is the only gold-containing mineral, it has been found that an 80% size of 30-40 microns or less is suitable for obtaining good gold extraction effects and acceptable leaching residence times.

In the case where gold is confined in pyrite, the pulverization size depends largely on the reactivity of the pyrite, and its reactivity has changed greatly as described above. In the case of the highly active pyrite, the particles used for the arsenite are used, but in the example of pyrite, which is more fire resistant, it is necessary to finely grind. In the case of more extreme refractory, the 80% particle size extends to ultrafine particles of 6-10 microns or less. In addition, the inventors recognize that the ultrafine grinding process has been developed to the extent that successfully operating multiple ultrafine mills in mines worldwide over the last decade.

Gold recovery

The leaching solution containing gold was passed through a column containing activated carbon to adsorb gold onto the activated carbon. The residence time for gold adsorption was 10-15 minutes, which is similar to that of the conventional method for cyanide systems. The weight of gold on carbon was usually 2-5% w / w due to the presence of concentrates with a relatively high gold content in the solution as a result of the use of graded concentrates with a relatively high gold content. Under such weight, gold was recovered by burning in the urinary tract to destroy carbon. If the load is lower, it is more economical to elute with cyanide and then reactivate the carbon.

Impurity Management

In addition to the main contaminants (eg arsenic, antimony, etc.), the presence of impurities (eg Cd, Mn, Mg, etc.) in the supplied concentrate does not have a detrimental effect on leaching or precipitation operations. Nevertheless, a method of managing impurities was used to prevent the accumulation of impurities in the treatment liquid over time. This was achieved via precipitation from the extract of the regenerated second copper solution and the purified brine was returned to the process. Importantly, IRGP produced no liquid effluent and removed all impurities as solid by-products.

Limestone was added to the extract to adjust the pH to 3.5 and the remaining iron and copper precipitated, then removed by filtration and recycled to the leaching step. Subsequently, impurities such as Cd, Mn and Mg were removed by forming an insoluble oxide by the addition of calcined lime at pH 9, recovered by filtration and discarded.

In terms of process equipment, IRGP is similar to the Biox method in that it uses atmospheric pressure, but the residence time is shorter, usually in the range of 6-20 hours. In the case of pyrite oxidation, higher leaching temperatures are used than in Biox, but when the concentrate supplied to the process is finely ground, oxygen plants (used for Pox) are usually finely ground to levels below 10 μm. You do not have to use. Materials constituting the process equipment were fiber reinforced plastics, rubber lined steel and titanium.

Ubiquitous  + Pyrite + carbon (double refractory ore)

The impact of carbon on the treatment of gold concentrates depends primarily on its grade and activity. When the carbon content is low, organic additives (blinding agents) are used to inhibit gold adsorption, or activated carbon is added in the leaching step to selectively adsorb gold (CIL; carbon treatment during leaching). Thus, in this case, the oxidation reaction of the arsenite is carried out as described above.

However, when the carbon content is high to exceed 3 to 5%, the effect of the inhibition method or CIL is greatly reduced, so-called "impregnation-extraction" of gold has increased. In this case, destroying carbon by the roasting method was the main treatment method conventionally performed. This is a relatively complex process since the extraction of gold from the calcined material formed is carried out under roasting conditions. In addition, the optimal condition for roasting pyrite is different from the ubiquitous roasting condition and thus requires a two-step roasting process.

Prior to roasting, the IRGP method can be used to selectively elute arsenic and sulfur to simplify subsequent roasting steps, in this case a simple one-step process. In addition, the removal of arsenic and sulfur reduces the burden of off-gas cleaning from the roaster operation, as the As ₂ O ₃ and SO ₂ are greatly reduced. Thus, this effect can significantly reduce the investment and operating costs of the roasting stage.

1st method and 2nd method

In the case of treating refractory sulfur-containing soils, in the first method according to the present invention, sulfide oxidation and dissolution of gold were simultaneously performed using a halide medium having a specific solution oxidation potential (so-called "all-in-one). " fair). In the second method according to the invention, halide media with different solution parameters are used to separate contaminants (e.g., arsenic) simultaneously with dissolution of some gold and to separate the contaminants before the sulfide is oxidized. Subsequently an additional amount of gold was recovered. Hereinafter, the first method and the second method according to the present invention will be described in detail.

1st way

With reference to the accompanying Figures 2 to 4 and Examples 1 to 3 will be described a first method.

In FIG. 2, a single refractory pyrite gold recovery method 10 is schematically illustrated. The precious metal concentrate 12 to be supplied to the method is prepared by mining, grinding, and flotation of sulfur-containing ore. The concentrate is usually gold-containing arsenite (which is double fire resistant if it contains a high content of carbon). The concentrate was ground to a very fine level, usually up to 10 μm, in a special ball mill. The pulverized concentrate was then fed to a first leaching step in the form of a arsenite leaching step 14.

In the arsenite leaching step 14, an acidic environment is maintained (preferably lower than pH 1 because leaching of arsenite is enhanced under low solution pH). The acid environment may be obtained only by recycling of the solution, or non-polluting acids (eg sulfuric acid or hydrochloric acid) may be added. The leaching solution Eh is usually maintained at 0.4 volts or more to promote oxidation of the arsenic iron component of sulfur-containing soils and solubilization of gold. The leaching temperature was maintained at about 80-95 ° C.

The leached soil was then fed to the second pyrite leaching step 16 where the pyrite was oxidized by adding an oxidizing agent (eg oxygen, air, chlorine, hydrogen peroxide, etc.) to raise the solution oxidation potential. In order to maintain the arsenic in the form precipitated in the second leaching step, it is necessary to add acid (eg sulfuric acid) or base (eg sodium carbonate) to maintain the pH of the solution above about 0.2.

The treatment solution was usually an aqueous solution of cupric chloride, with a chloride concentration of 8 mol / liter. In both the arsenic and pyrite leaching steps, the second copper ions oxidized the sulfur containing soil and reduced to the first copper ions (Scheme 2 and 8 above). In addition, the second copper ions were regenerated in an acidic oxidation environment (Scheme 3 and 9). Thus, in such a method, copper acts as an electron transfer agent present in Cu ²⁺ / Cu ⁺ pairs. Other reagents may perform this action, for example, iron, cobalt, manganese, vanadium and the like.

If the carbon content of sulfur-containing soils is high (eg 3-5 wt% or less), the gold (or other precious metals) leached into the solution by adding the shielding surfactant in steps (14) and (16) is carbon phase. Adsorption to can be prevented. Surfactants are usually organic blinding agents such as kerosene, phenol ether and the like. Alternatively, activated carbon may be added to preferentially adsorb gold and subsequently remove it.

In the first method, in the ubiquitous leaching step 14, the inventors found that at a controlled pH at which pH 1 or less but arsenic is solubilized, at a relatively moderate oxidation potential of about 0.7-0.8 volts (based on SHE), It has been found that when using low temperatures (80-95 ° C.), gold can be solubilized by leaching sulfur-containing soils without oxidizing pyrite sulfides to sulfates.

The oxidation conditions used in the pyrite leaching step 16 are more severe than the conditions used in the arsenic leaching step 14. Here, the oxidant was sprayed onto the solution to oxidize the oxidation potential Eh to about 0.85 volts. In addition, the temperature of the solution in the second leaching step needs to be raised to about 90-105 ° C. Likewise, in the second leaching step the pH of the solution according to the first method is adjusted to at least 1 but to a pH at which arsenic is solubilized.

Since acid is consumed in the second leaching stage (ie, Cu (II) is regenerated), it is necessary to add acid to the solution of leaching stage 16 periodically or continuously, such as sulfuric acid, hydrochloric acid or Other acids do not chemically interfere with the process. However, the amount of acid replenishment depends on whether enough sulfuric acid is produced by leaching of pyrite. In addition, the pH was adjusted by adding calcium carbonate in the leaching step 16 to prevent solubilization of arsenic.

In the leaching step 16, the sulfide is oxidized to sulphate, iron is eluted into the solution in the form of Fe (III) (Scheme 1), and gold (or other precious metals) usually remaining in pyrite is solubilized. The inventors have unexpectedly found that an oxidation potential in the 850 mV (SHE basis) region can be obtained under the influence of Cu ²⁺ and Fe ³⁺ in a halide solution in the presence of an acid. This potential was sufficient to dissolve the gold through the formation of the gold-chloride complex in the 8M Cl ⁻ medium used.

The solid slurry obtained in the pyrite leaching step 16 was fed to the solid-liquid separation step 18, where the solid was filtered out of the solution using a known filtration apparatus. The obtained filtrate 20 was fed to an in-line noble metal recovery step 22, and the filtered solid was disposed of as tail. Supplemental water was added in step 18 to make up for the water lost with the residue.

The metal recovery step 22 comprises one or more columns filled with activated carbon, for example fluidized bed arrangement, through which the solution is passed upstream. Gold (or other precious metals) solubilized in solution is adsorbed to activated carbon, while upstream liquid stream 26 exits the column and is recycled to leaching step 14. The activated carbon containing gold is then removed or periodically treated to pass as a gold product stream 28 for the gold recovery process (eg, carbon product is calcined or the carbon column is eluted with cyanide solution).

The upstream liquid stream 26 is recycled to the leaching step 14 via an iron precipitation step in the form of a solution conditioning step 30. In step 30, sulfur and iron are removed from the process by precipitating the soluble ferric sulfate obtained in the pyrite oxidation step 16 and adding limestone and calcium carbonate to form hematite and gypsum (Scheme 6). The amount of limestone added is adjusted to maintain the iron concentration in the solution at approximately 2 g / l to prevent precipitation of copper in the second copper form. The hematite / gypsum slurry is filtered and the residue washed, then disposed of with waste. The solution is then recycled to step 14 above.

To prevent contaminants from accumulating throughout the process, part 32 of stream 26 is recycled to extract circuit 34 to separate contaminants such as Mn, Cd, Ni, Co, etc. (e.g., For example by raising the extract solution pH to precipitate in a controlled manner).

First Method Example

In the following, the optimum process route will be described through the preferred embodiment of the first method according to the present invention.

Example 1

As a preliminary evaluation, the method of extracting gold from the first concentrate was carried out in three different steps: arsenite leaching, pyrite leaching (1) and pyrite leaching (2). The following experimental test report details the procedures and results of the three steps. CON1 01 refers to arsenite leaching and pyrite oxidation (1), and CON1 02 refers to pyrite oxidation (2). The first concentrate was ground to P80 = 30μ and treated with an As leaching step followed by a pyrite oxidation step.

purpose

The purpose of this embodiment is to apply IRGP to a single refractory Au concentrate. Ore samples were provided to metallurgical laboratories for grinding and concentration.

step

The experiment was conducted in two parts and in a 7.5 L titanium insulated tank. The first portion was an As leaching portion, using a conventional mixer. The second portion, as a pyrite oxide portion, used a flat blade turbine and sparging device.

First part: As leaching

In a 7.5 L titanium reactor with a "propeller" stirrer, 3.5 L of neutral brine with a pH <0.5 was prepared using 200 g / l NaCl, 50 g / l CaCl ₂ . In addition, a "boost" solution having a pH <0.5 was prepared using 200 g / l NaCl, 50 g / l CaCl ₂ and 75 g / l CuCl ₂ form Cu. If necessary, several grams of copper dendrites were added to adjust the Eh in the range from 580 to 600 mV. The synergistic solution was maintained at a temperature of 80 ° C.

After heating the leaching reactor to 105 ° C., 300 grams of anhydrous concentrate was added to the brine. After 15 minutes of necessity, concentrated hydrochloric acid was added to the suspension to adjust the pH to <0.5 and samples were taken at t = 0. It was confirmed that all acid was added (time, addition volume, volume in the leaching tank).

Eh and pH were measured and monitored to ensure that Eh did not exceed 530 mV while slowly adding the synergistic solution at a rate of 2.5 l / hour. Solution samples were taken every 30 minutes for As, Fe, Cu analysis. Eh and pH were monitored every 30 minutes.

When Eh reached 530 mV and stabilized, the As leaching step was considered complete. The cake formed was washed twice with hot brine (50 gpl NaCl, pH <1.0) then washed with hot water until the filtrate became clear. The cake was then dried in the oven overnight. The cake was analyzed for S _(T) , S _(E) , As, Fe, Au and C.

Second part: pyrite oxidation

The 7.5 L reactors were equipped with flat blade turbines and titanium sparging tubes. 10 liters of saline solution adjusted to pH <0.5 using 200 g / l NaCl, 50 g / l CaCl ₂ and 75 g / l CuCl ₂ form Cu and adding 8.8 mol of concentrated hydrochloric acid were added to the leaching tank. Prepared. The solution was heated to 105 ° C., samples were taken at t = 0 and then the dry cake produced in the first leaching As leaching portion was placed in a tank. After 15 minutes, a solution sample was taken to measure Eh and pH. Technical grade HCl was added as needed to adjust the pH to <0.5.

Oxygen was injected at a rate of 2 l / min, Eh and pH were monitored every 30 minutes, and samples were taken for Fe, As, Cu analysis every hour. The leaching step was considered complete when Eh was stable above 600 mV for 3 hours and Fe did not change in solution. The slurry was filtered. The cake formed was washed twice with hot brine (50 gpl NaCl, pH <1.0) and then with hot water until the filtrate became clear. The cake was then dried in the oven overnight. The cake was analyzed for S _(T) , S _(E) , As, Fe, Au and C.

Example 2

The pyrite oxidation was further performed on the concentrate residue obtained in Example 1 as follows.

purpose

Analysis of the data and residues obtained from Example 1 showed that the pyrite oxidation was not complete at the end of the experiment. The procedure of this example using an improved brine composition is intended to increase the effectiveness of the Au extraction method using oxygen to oxidize pyrite.

result

The second pyrite oxidation step showed improved metal extraction results as shown in Table 1 below (independent analysis results).

element Oxidation reaction 1 Oxidation reaction 2 As 79.6% 92.4% Fe 72.2% 97.1% Au 68.7% 93.3%

step

The 7.5 L reactors were equipped with flat blade turbines and titanium sparging tubes. Five liters of saline solution, adjusted to pH <0.5, were prepared in a leaching tank using 100 g / l NaCl, 250 g / l CaCl ₂ and 100 g / l CuCl ₂ form Cu and adding concentrated hydrochloric acid. The solution was heated to 105 ° C., samples were taken at t = 0 and then the dry cake produced in the first leaching As leaching portion was placed in a tank. As leaching / pyrite oxidation product was introduced into the tank. After 15 minutes, a solution sample was taken to measure Eh and pH. Concentrated HCl was added as needed to adjust pH to <0.5.

Oxygen was injected at a rate of 2 l / min, Eh and pH were monitored every 30 minutes, and samples were taken for Fe, As, Cu analysis every hour. When Eh was stable above 600 mV for 3 hours and Fe did not change in solution, oxygen inlet was stopped and Eh was monitored. When the Eh stayed above 600 mV, the leaching step was considered complete.

The slurry was filtered. The cake formed was washed twice with hot brine (50 gpl NaCl, pH <1.0) and then with hot water until the filtrate became clear. The cake was then dried in the oven overnight. The cake was analyzed for S _(T) , S _(E) , As, Fe, Au and C.

result

The results obtained through the experiments of Examples 1 and 2 are shown in Table 3 below.

     Duration T Eh (mV) pH Fe Accumulation As Accumulation Fe As (hours) (° C) (g) (g) (g / l) (g / l) Ars Pyrite oxidation 1 P

Gold extraction results are shown in FIG. 3.

Example 3

In this experiment, gold was extracted from the second concentrate in three successive stages: Step 1 (ferrous and pyrite leaching), Step 2 (pyrite leaching with oxygen) and step 3 (pyrite using chlorine). Leaching).

purpose

After the reconnaissance experiment for As leaching, this experimental procedure is intended to treat the second concentrate in an "all-in-one" manner using 250 g / l CaCl ₂ and 100 g / l saline Cu. Solid weight was set at 200 g / l.

result

The chlorine oxidation reaction is as shown in Table 4 below.

Au extraction (accumulated amount) Air + oxygen 59% Goat 87% Sum 95%

step

The procedure of this example was carried out in a 7.5 liter reactor with a turbine stirrer. 5 l of brine with the following composition were prepared: pH was adjusted to <0.5 by addition of 100 g / l NaCl, 250 g / l CaCl ₂ and 100 g / l Cu and concentrated hydrochloric acid.

First part: Ubiquitous  Leaching

The brine was heated to 90 ° C. with the stirrer rpm at 90%. Solution samples were taken for later reference. 1000 g equivalents of dried “as received” concentrate (P80 ca. 37 μ) were added to the brine. Eh and pH were recorded 15 minutes after the sample was taken at t = 0.

Air was injected into the reactor at a rate of 2 l / min. Eh and pH were monitored every 30 minutes and solution samples were taken for As and Fe analysis. When Eh and Fe in solution became stable, the air flow was shut off. When Eh fell 20 mV or more, the introduction of air was resumed for 2 hours. If the Eh did not fall above 20 mV, about 100 g of solid sample was taken and fed with air replaced with oxygen.

Second part: pyrite oxidation

The temperature was raised to 105 ° C. Sampling and measurement frequency were changed at 1 hour intervals. Oxygen was injected under the turbine stirrer at a rate of 2 l / min. When Eh and Fe in the solution became stable, the oxygen supply was stopped. If Eh fell by more than 20 mV, oxygen injection resumed for 2 hours. If Eh did not fall above 20 mV, the experiment was considered complete.

The suspension was filtered and the cake washed twice with acidic brine and then with hot water until a clear filtrate was obtained. The washed cake was dried and weighed. The residue was analyzed for As, Fe, Cu, element S, total S and Au. Final solution samples were also analyzed for Au.

Third part: pyrite chlorination

To improve the Au extraction effect, the experiment was extended to pyrite chlorination using hypochlorite as the chlorine source. The residue obtained from pyrite oxidation using oxygen was added to 4 l of brine having the same composition as above. The temperature was raised to 100 ° C. or higher and 50 g hypochlorite was added every 30 minutes. Fe concentration was monitored. When the Fe concentration did not change even after hypochlorous acid addition and Eh was stable, the experiment was considered complete.

 The suspension was filtered and the cake washed twice with acidic brine and then with hot water until a clear filtrate was obtained. The washed cake was dried and weighed. The residue was analyzed for As, Fe, Cu, element S, total S and Au. Final solution samples were also analyzed for Au.

result

Duration Eh pH Fe Total Fe (hours) (mV) (g / l) (g) air Oxygen Hypochlorite

Gold extraction results are shown in FIG. 4.

2nd way

Before describing the second method in detail with reference to the embodiment, the second method will be described in general with reference to FIG. 5.

In FIG. 5, the precious metal concentrate 10 to be supplied to the process is obtained by mining, grinding and flotation of sulfur-containing ores. In a second method, the concentrate can be a gold-containing arsenite, which has a high carbon content (eg 2-20% by weight carbon) or a low carbon content or carbon-free (eg 2% by weight or less). . The concentrate is ground in a ball mill 12 and then fed to the contaminant oxidation step in the form of an arsenic leaching step 14.

Preferred arsenic leaching steps are described in more detail with reference to FIGS. 6 and 7 and Examples 11 and 12 below. Leaching can be carried out in a single step (for example using one or more treatment units, vessels or tanks), but usually in a multistage (two step) process. Each stage may comprise one or more processing units, vessels or tanks configured co-current or counter-current and upstream and downstream (in the form of a notification).

In any case, in the leaching step 14, a high acidic environment is maintained (preferably lower than pH 1 because the pH of the solution is preferably low to leach arsenic from the arsenite). The acid environment may be obtained only by oxidizing sulfur-containing soils (eg when sulfur in sulfur-containing soils are oxidized to sulfates in solution), or non-contaminating acids may be added (eg sulfuric acid or hydrochloric acid).

In addition, in the second method, the Eh of the leach solution is maintained at 0.4 volts or more (see FIG. 8) to solubilize contaminants (eg arsenic). As described below in connection with FIGS. 6 and 7 and the embodiment, the leaching process includes two steps. In the first leaching step, the Eh of the solution is carefully controlled to oxidize and solubilize to a +3 oxidation state rather than a relatively poorly soluble +5 state, so as not to substantially oxidize pyrite in the arsenite. In the second leaching step, pyrite is oxidized by adding an oxidizing agent (e.g., oxygen, air, chlorine, hydrogen peroxide, etc.) to raise the oxidation potential of the solution (at the same time, the remaining arsenic is oxidized to a +5 oxidation state). In the second leaching step of the second method, As (V) is maintained in the solution or precipitated by adding an acid (e.g. sulfuric acid) in a controlled manner to lower the pH of the solution to a point sufficient to solubilize the arsenic. It can be kept in form and discharged from the process with pyrite residues.

Here too, the treatment solution is an aqueous cupric chloride solution, preferably having a chloride concentration of 7-8 mol / liter. Copper also acts as a leaching reagent and as an electron transfer agent.

In addition, in the case where the carbon content of the sulfur-containing soil is high (for example, 2% by weight or more), a shielding surfactant may be added in step 14 to prevent the noble metal leaching into the solution from adsorbing onto the carbon.

At an acid pH of 1 or less, under controlled Eh in the range of 0.45 to 1.25 volts, most suitably at about 0.5 volts Eh, arsenic is oxidized without oxidizing the sulfides of pyrite into sulfates that interfere with the properties of the solution. In solution, it can be eluted, preferably in the +3 oxidation state.

In the first leaching step, the ubiquitous concentrate is leached for a predetermined period of time (as described in the following examples) until a predetermined amount of arsenic is eluted from the ubiquitous (typically about 85 of the total amount in the first leaching step). % Is leached, in the second leaching stage an additional 10% of the total amount is leached). In any case, the amount leached is determined by the allowable residual concentration in the leached arsenite, assuming that the leached arsenic is separated and subsequently treated or disposed of by conventional smelting or roasting techniques (described below). The term "refined arsenite" or "refined sulfur-containing soil" is interpreted in this sense.

Thus, in the second method, the pH and Eh of the solution are adjusted such that arsenic and polyvalent species Cu (II) (oxidize and elute arsenic from the soil) are present in the solution but do not precipitate out of the solution in the first leaching step. .

In addition, the operating conditions of the method are adjusted such that arsenic remains in the solution in the solid / liquid separation step (separating the refined arsenite solid from the solution) until the solution is fed to the arsenic precipitation step. In FIG. 5 this step is shown schematically as a concentration step 16. In the process shown in Figures 6 and 7, the concentration step is used after the first leaching step. In the concentration step 16, the refined ubiquitous solid is agglomerated (ie, added flocculant), and the solid is extracted as a downstream stream 18, and arsenic and precious metals contained in the supernatant are removed from the concentration step by an upstream stream ( 20). In FIG. 5, the downstream stream or slurry 18 is then fed to a solid-liquid separation step 22, in which the solid is typically filtered out of solution using a known filtration device.

The resulting filtrate 24 is returned to the upstream stream 20, while the filtered solid (i.e., refined arsenite) 26 is subjected to a conventional roasting step 28 and a conventional cyanide leaching step 30. To recover the precious metals remaining as gold product (32).

Depending on the degree of leaching of the precious metal in the arsenic leaching step 14, gold (and other precious metals) may be treated as a combined liquid stream 34 (combination of streams 20 and 24) and recovered through inline precious metal recovery step 36. have. The recovery step includes one or more columns filled with activated carbon and passes the solution upward through activated carbon arranged in a fluidized bed. Gold (or other precious metals) dissolved in the solution is adsorbed onto the carbon, while arsenic dissolved in the solution passes the pair as an upstream liquid stream 38. The activated carbon containing gold is then periodically removed and fed as a gold product stream 40 (along with the gold product stream 32) to a gold recovery step (eg calcining or elution of the carbon product).

From the metal recovery step 36, a solution (including dissolved arsenic) 38 is fed to the contaminant precipitation step in the form of an arsenic precipitation step 42. Step 42 is usually carried out at pH 1.5-3. In step 42, an oxidizing agent (e.g., air, oxygen, chlorine, etc.) is introduced into the solution (e.g. sparging) to raise the oxidation potential (Eh) of the solution, thereby causing dissolved arsenic to precipitate, usually insoluble iron arsenate. Allow to form a precipitate (ie FeAsO ₄ or scorodite). If the contaminants include, for example, antimony, the contaminants will form iron antimonate in an insoluble form. As precipitation of contaminants is formed, acid is produced, so that alkalis can be added to consume the acid to maintain optimal solution pH and Eh. Typically, alkali is an alkali salt such as calcium carbonate or calcium oxide, and the use of alkali also provides another advantage of allowing the precipitation of sulfates in the treatment liquid.

In the second method, the amount of oxidant and alkali is adjusted to maintain optimal pH and Eh levels until all contaminants have precipitated in the contaminant precipitation step 42. The solution may then be recycled to the leaching step after the contaminant precipitation step by returning the pH and Eh levels of the solution to the corresponding level in leaching step 14 as needed.

In addition, in the precipitation step of the second method, the oxidant oxidizes the copper in the first copper form to the copper in the second copper form, thereby regenerating the chemical species to enable recycling and reuse. Thus, by regulating the Eh and pH of the solution, it promotes the reoxidation of the polyvalent species and at the same time ensures that the polyvalent species always remain in the solution, alternating copper to +1 and +2 oxidation states throughout the process. It is an advantageous feature to act as an electron transfer agent and a precipitate upon leaching. Regeneration of multivalent species is economically beneficial, simplifies the process and perfects mass balance.

After arsenic precipitates, the arsenic precipitate is separated from the treatment liquid in a solid / liquid separation step. In FIG. 5, this precipitate is shown as an additional concentration step 44 to produce a solid (arsenic precipitation) downstream product 46, which is then fed to a solid-liquid separation step 48. Supernatant upstream stream 50 is withdrawn from concentration step 44. In the solid-liquid separation step 48, arsenic waste 52 is typically produced by filtration off arsenic precipitate using a filtration device. The filtrate is returned as a stream 54 of liquid phase to the upstream stream 50. The combined liquid stream 56 is then fed to a further noble metal recovery step 58, such as an activated carbon column, to recover the metals not recovered in step 36. Alternatively, step 58 may be used instead of step 36. The obtained precious metal and activated carbon stream 60 is combined with the other precious metal recovery streams 40, 32 while the solution upstream stream 62 is recycled to the arsenic leaching stage 14 to form a closed recovery treatment loop.

In order to deal with the accumulation of contaminants throughout the process, a portion of the recycle stream 62 is recycled to the extraction circuit 64 so that contaminants not recovered in the arsenic precipitation step and, in some cases, other contaminants, for example For example, Mn, Cd, Ni, Co and the like can be separated.

The foregoing has outlined the second method, and an embodiment of the preferred second method will be described below with reference to FIGS. 6 and 7.

6 shows a process diagram of a method for treating a single refractory sulfur-containing soil. In FIG. 6, steps similar to those of FIG. 5 are indicated using similar reference numerals. In a manner similar to the method of FIG. 5, ubiquitous concentrates containing gold and low or high carbon content (ie, single refractory soils) are prepared and ground (10, 12). The ground concentrate is then fed to a preferred leaching process. The preferred leaching process consists of two stages: the first arsenic (FeAsS) leaching step 70 and the second pyrite (FeS ₂ ) leaching step 72.

Ubirite concentrate is fed to the first leaching step 70, where the leaching conditions are adjusted to oxidize only arsenic in the concentrate to leach into solution and not to oxidize the pyrite component of the concentrate. Here, the leaching conditions of the first leaching step 70 are such that the oxidation potential Eh is about 0.5 volts, the pH of the solution is 1 or less, and the temperature of the solution is maintained at about 105 ° C., provided that the first leaching step 70 is performed. May be carried out at a temperature in the range from 80 ° C. to 105 ° C.). We have found that these conditions are optimal for leaching arsenic into solution. As explained in Example 11 below, after leaching for about 6 hours, about 85% of the total arsenic component of the arsenite concentrate was leached into the solution.

When a predetermined amount of arsenic has leached into the solution, the solution and arsenic-free arsenic iron are fed to the concentration step 16 in a manner similar to the method shown in FIG. The refined arsenite solid is agglomerated and extracted as downstream stream 18, while the arsenic-impregnated supernatant is discharged from the concentrator to the upstream stream 20 and fed to the arsenic precipitation stage 42.

In a second method, the refined ubiquitous solid stream 18 is then fed to a second leaching step 72 for leaching pyrite. The oxidation conditions in the second leaching stage are more severe than the conditions of the first leaching stage. Here, an oxidant such as oxygen is sparged in the solution to increase the oxidation potential Eh to 0.6 volts or more, usually 0.8 volts or more. In addition, the temperature of the solution in the second leaching step is maintained at about 105 ℃. In the second leaching step the pH of the solution is still maintained at pH 1 or below.

Since acid is consumed in the second leaching stage (ie Cu (II) and Fe (III) are reduced to Cu (I) and Fe (II), respectively), chemically sensitive to sulfuric acid, hydrochloric acid or the process periodically or continuously. It may be necessary to supply acids that do not interfere. However, the need for supplemental acid depends on whether sufficient sulfuric acid is produced by elution of pyrite. Keeping the pH low in the second leaching step also helps to solubilize As (V) as needed.

In the second leaching step, the sulfur-containing soil is oxidized with sulfate, iron is eluted into the solution as Fe (III), and a portion of the arsenic remaining in the arsenite is eluted into the solution. The inventors have found that 10% of the total arsenic component is leached into the solution so that the arsenic finally remaining after the leaching process is less than 5% of the total amount of the concentrate raw material. This corresponds to arsenic concentrations low enough to safely dispose of residues from the process.

The leached product obtained in the second leaching stage is fed to the solid-liquid separation stage 22 similar to that shown in FIG. 5 as stream 74, so that the remaining solids are filtered out of solution and the filtrate stream 24 is upstream. Return to stream 22 and combine with it to feed arsenic precipitation step 42. The solid residue filtered in step 22 is then treated as stream 76 with debris in the form of a filtered solid or slurry. As an alternative, the solids may be further treated for residual metal recovery. Water may be added to step 22 to maintain the moisture content in the process and / or to supplement the water lost with the process residue.

Since gold or other precious metals of single refractory arsenite and pyrite do not bind carbon to a significant extent, both in the first leaching step and in the second leaching step can be eluted into solution and thus recovered in the process circuit.

In the method shown in FIG. 6, in the arsenic precipitation step 42, the pH of the solution is adjusted to about 2 to 3 (eg by adding calcium carbonate), and an oxidant such as air or oxygen is added to the solution. The arsenic is then oxidized from the soluble +3 form to the insoluble +5 form. Since iron (III) is present in the solution through oxidation of pyrite in the second leaching step, arsenic is precipitated as scorodite (FeAsO ₄ ). As another advantage, since the sulfide is oxidized to sulfate in the second leaching step, calcium carbonate can be added to raise the solution pH of the arsenic precipitation step, and the sulfate can be precipitated as calcium sulfate.

Arsenic / solid precipitate is then fed along with the treatment solution to the solid-liquid separation stage 48 as stream 78, whereby the reverse solids are filtered out of solution. Solid residue stream 80 typically comprises FeAsO ₄ , Fe ₂ O _3, and CaSO ₄ in a form suitable for disposal (eg, landfill). Since the solid can be removed as a slurry, make-up water can be added to step 48. Thus, arsenic, iron and sulfur are advantageously recovered from the original ubiquitous concentrate in a form that is easy to dispose of.

Since the conditions of the arsenic precipitation step do not affect the precious metal leached into the solution in the leaching step, the separated solution 56 can now be supplied to the precious metal recovery step 58 in a manner similar to the method shown in FIG. 5. Step 58 includes one or more columns containing activated carbon, on which the precious metal, in particular gold, is adsorbed. In addition, the gold product stream 60 is periodically removed from step 58 to recover the gold (collected by eluting the calcined or adsorbed carbon).

As in the method shown in FIG. 5, the solution upstream 62 from step 58 is recycled to the leaching process and a portion of the regenerated stream is discharged to the extracting circuit 64 to contaminants that may accumulate in the process. By separating off, contaminant byproduct stream 82 is produced.

In a second method, the recycle stream 62 of the solution is split to produce a first leach stage recycle component 84 and a second leach stage recycle component 86. In the case of the copper chloride treatment liquid, copper in the +2 oxidation state is recycled to each leaching step to participate in leaching of arsenite in the first leaching step and leaching of pyrite in the second leaching step.

Referring to FIG. 7, a process diagram of a second method for producing double refractory sulfur-containing soils is shown. In Fig. 7, processing steps similar to those of the method shown in Figs. 5 and 6 are shown using similar reference numerals. In addition, since the upper half (i.e., dotted line) of the process diagram shown in FIG. 7 is the same as the process step of FIG. 6, this part is not described again.

In double refractory arsenite, precious metals (such as gold) are usually associated with carbon, so gold does not readily leach into solution in the first and second leaching steps. Thus, solid stream 76 comprises a carbon / gold component combined with a solid residue. However, because the elution method has substantially reduced arsenic, iron, sulfur and other contaminants to acceptable low levels, it is well suited for roasting or smelting solid residues from the leaching process in the roasting step 28.

In the roasting step 28, the fuel and air are roasted in a conventional manner with the solid residue 76 to produce a product stream 90, which is then subjected to a gold leaching step 30 in a known manner. Supplies). Leaching of gold is usually done by oxidizing the calcined solid to chlorine gas or cyanide (preferably because chlorine gas has a lower toxicity than cyanide). In the second method, a portion of the solution 92 from the arsenic precipitation step 42 can be recovered, creating an economic effect for all processes in the gold leaching step 30.

The exhaust stream 94 of the roasting stage 28 (usually including carbon dioxide, sulfur dioxide and other So _x gases) is fed to the primary gas scrubbing stage 96. The primary gas scrubbing step typically comprises one or more scrubbers, where water and optionally recycle scrubbing water are brought into contact with the gas stream 94. It may be advantageous to filter the dust in the gas stream 98. Such dust may include gold chloride (AuCl ₃ ) and arsenic oxide (AS ₂ O ₃ ). These dusts, along with other particulates, can be fed back to the arsenic precipitation step 42 in solid or solution form as stream 98 to recover additional amounts of arsenic and gold.

Residual gas discharged from the primary gas scrubbing stage 96 is fed to the secondary gas scrubbing stage 102 as a stream 100, which typically contacts the calcium carbonate in solution with the SO _x containing gas. Gas scrubber. Thus, the product stream obtained from the secondary gas scrubbing step 102 typically includes calcium sulfate and calcium sulfite.

The leached product 106 containing the dissolved gold obtained in the gold leaching step 30 is now fed to the solid-liquid separation step 108 to separate the gold containing solution from the solid residue. The solid residue stream 110 is disposed of as waste, while the supernatant containing gold is fed to a gold recovery step 104, typically a column containing activated carbon. Carbon and adsorbed gold are periodically removed as stream 116 to recover gold. On the other hand, the solution 118 which is almost free of gold is recycled to the leaching / arsenic removal step and combined with the stream 34.

Second Method Example

Hereinafter, an optimal flow path for the second method will be described according to a preferred embodiment for the second method. In the examples described below, the high refractory arsenite concentrates obtained from Bakyrchik, Kazakhstan were treated. The purpose of this example is to develop a process that enables the processing of all samples of ubiquitous iron provided from the Bakhkirch mine.

Example 4: Concentrate Characterization

Way:

6 kg of concentrate was treated by ultrafine grinding. Particle size of the concentrate is P ₁₀₀ 20 microns as the value.

product Laser micron Weight percent passed 20 100 18 99 15 96 12 89 10 81 8 69 6 50 5 42 3 15

Concentrates with a P ₁₀₀ of 20 microns were obtained in the form of three cakes, and the moisture content of each cake was measured and the average value was used as the moisture content for the concentrate.

Cake 1

Wet state sample + paper: 113.84 g

Dry sample + paper: 85.68 g

Paper: 4.83 g

Dry sample: 80.85 g

% Moisture content: 25.8%

Cake 2

Wet state sample + paper: 88.35 g

Dry sample + paper: 66.65 g

Paper: 4.83 g

Dry sample: 62.02 g

Moisture Content (%): 25.9%

Cake 3

Wet Sample + Paper: 86.41 g

Dry sample + paper: 68.79 g

Paper: 4.85 g

Dry sample: 63.94 g

% Moisture content: 21.6%

The average water content was measured at 24.4%. From this, it can be seen that 100 g of the dry concentrate corresponds to 132.3 g of the wet concentrate sample, calculated according to the calculation.

Example 5

Leaching by Oxidation Reaction

Tests were performed on regrind samples with P ₁₀₀ = 20 microns to provide initial evidence that arsenic leached through the oxidation reaction. Bakirch ore concentrates are known to contain arsenic in the form of arsenite. The reaction is designed to see if this arsenic can be converted to soluble (and thus selectively removable) when using copper in the form of a second copper as oxidant.

Way

1 L of a solution containing 80 g / L Cu ²⁺ (205.13 g as CuCl ₂ ), 100 g / L CaCl ₂ , 200 g / L NaCl and 30 g / L NaBr was prepared. 140 g of wet concentrate (about 24% water content, ie 105.8 g of dry concentrate) were added to the solution and the resulting slurry was stirred at 105 ° C. pH, Eh and Fe and Cu contents were measured over 4 hours.

The solid was then filtered using a Buchner apparatus and the filtrate was preserved for analysis. The solid cake was washed with low pH brine (about 0.5 L, 280 g / L, pH 0.3) and the resulting wet cake was weighed, dried in an oven and weighed again. The dried solid was saved for later analysis.

Results and Discussion

The pH, Eh and Fe and Cu contents recorded over time are summarized in Table 5 below.

Minutes pH Eh (mV) Fe (g / L) Cu ^{total amount} (g / L) 0 30 1.45 740 2.4 70.6 60 0.5 508 2.5 68 90 0.5 507 2.5 63 150 0.5 495 2.65 64 210 0.35 502 2.65 61 270 0.35 495 2.68 64 330 0.35 485 2.66 65

Residue analysis showed that the As concentration was 0.66%. In view of the calculated mass loss of 6.5%, the leaching efficiency of As was 82.3%.

The reaction seemed to go fast. Eh and pH were markedly lowered within the first hour of the reaction progression. After this time, the reaction was stabilized and showed no progress.

Example 6

Leaching by Oxidation Reaction

The purpose of this example is to find out whether the new treatment liquid further leaches iron / arsenic from the previously leached soil. It is speculated that the treatment of the solids obtained from the pre-leaching step further removes arsenite. A preliminary leaching step was prepared by preparing a fresh solution of the primary treatment liquor and using the soil subjected to the leaching step as a solid feedstock.

Way

1 L of a solution containing 80 g / L Cu ²⁺ (102.55 g as CuCl ₂ ), 100 g / L CaCl ₂ , 200 g / L NaCl and 30 g / L NaBr was prepared. 30 g of the leached treated concentrate obtained from the pre-oxidation step was added to the solution and the resulting slurry was stirred at 105 ° C. pH and Eh were measured over 4 hours. The solid was then filtered using a Buchner apparatus and the filtrate was saved for analysis. The solid cake was washed with low pH brine (about 0.5 L, 280 g / L, pH 0.3) and the resulting wet cake was weighed, dried in an oven and weighed again. The dried solid was saved for later analysis.

Samples were taken from the solids obtained from the pre-reaction step and the solids obtained as above and the original concentrate was immersed using Aqua-regia / percholic acid. The solution was then analyzed for arsenic using ICP.

Results and Discussion

The pH and Eh recorded over time are summarized in Table 8 below.

Minutes pH Eh (mV) 0 1.32 741 0 1.2 615 30 0.55 588 60 0.31 583 90 0.29 580 120 - 579 150 0.31 569 180 0.3 574 210 0.29 574 240 0.32 572

Wet Cake + Paper + Filter: 72.5 g

Dry Cake + Paper + Filter: 40.24 g

Paper + filter paper: 11.5 g

Obtained dry cake: 28.74 g

The ICP analysis of arsenic in the recovered solids is summarized in Table 9 below.

As content extraction (weight%) (As% in concentrate) Bachkirch Concentrate 3.49 0 Leaching stage 1 0.66 82.3 Leaching stage 2 0.42 88.9

As observed in the preleaching step, the reaction appeared to proceed rapidly, and stabilized after 1 hour. Likewise, the Eh and pH decreased markedly with the decrease in mass of recovered solids as compared to the mass of solids fed into the solution. This suggests that there is still extractable soil in the residue obtained from the first leaching step. Analysis of the arsenic content of the feedstock and the solid residue obtained from the two leaching stages revealed that the arsenic content of the recovered solids gradually decreased. This result suggests that the method of the present invention can be reconfigured to selectively leach arsenic contained in bakirchixane concentrate.

Example 7

Leaching by Oxidation Reaction

The purpose of this example is to reconstruct the conditions used to leach arsenic from bakirchixane ore. Although successful in eluting about 65% of the arsenic contained in the bakirchixane ore, the process of the present invention was reconstituted to obtain higher leaching performance. The focus of the method of the invention is in two areas. That is, the first is to simply configure the leaching solution, and the second is to investigate the effect of temperature and pH changes on the leaching effect by performing the reaction at various temperatures and starting pH.

Way

5 L of a solution containing 80 g / L Cu ²⁺ (1025.64 g as CuCl ₂ ), 150 g / L CaCl ₂ (750 g) and 150 g / L NaCl (750 g) was prepared and heated at 80 ° C. Subsequently, the solution was divided into 1.5 L portions, and each solution was used for leaching by oxidation under 142.86 g equivalents of wet concentrate (about 24% water content, ie 108 g in terms of dry concentrate) under different conditions. It was.

Oxidation leaching treatment solution 1: A leaching step was performed at 80 ° C

Oxidation leaching treatment solution 2: A leaching step was performed at 100 ° C

Oxidation Leach Treatment Solution 3: Leaching step at 80 ° C., initial pH <0.4, Eh> 550 mV

The pH and Eh of the solutions were measured over two and a half hours. Samples from each test were taken at timed intervals and analyzed for iron and copper content.

The solid was then filtered using a Buchner apparatus and the filtrate was saved for analysis. The solid cake was washed with low pH brine (about 0.5 L, 280 g / L, pH 0.3) and the resulting wet cake was weighed, dried in an oven and weighed again. The dried solid was saved for later analysis.

Samples were taken from the solids obtained from each reaction and analyzed for arsenic, copper and iron using ICP by dipping the original concentrate.

Results and Discussion

Oxidation Leach Treatment 1 (80 ° C) Solution Analysis Minutes pH Eh (mV) Cu (g / L) Fe (g / L) As (ppm) ICP measurement Reference 0 2.15 720 Solid addition 0 1.58 568 71 - - 30 0.95 535 77 0.79 824 60 0.75 525 77 1.14 1033 90 0.70 520 75 1.28 1152 120 0.70 520 75 1.41 1216 150 0.70 516 74 1.53 1308

Wet Cake + Paper + Filter: 173.24 g

Dry Cake + Paper + Filter: 105.48 g

Paper + filter paper: 11.5 g

Obtained dry cake: 93.98 g

Oxidation Leach Treatment 2 (100 ° C) Solution Analysis Minutes pH Eh (mV) Cu (g / L) Fe (g / L) As (ppm) ICP measurement Reference 0 1.88 735 Solid addition 0 1.5 561 71 - - 30 1.1 525 79 1.75 1473 60 1.05 528 79 1.85 1532 90 0.98 529 80 2.0 1636 120 0.94 525 84 2.0 1678 150 0.89 528 85 2.14 1761

Wet Cake + Paper + Filter: 170.3 g

Dry Cake + Paper + Filter: 113.32 g

Paper + filter paper: 11.5 g

Obtained dry cake: 101.82 g

Oxidation Leach Treatment 3 (80 ° C, Low pH) Solution Analysis Minutes pH Eh (mV) Cu (g / L) Fe (g / L) As (ppm) ICP measurement Reference 0 0.35 712 Solid addition 0 0.8 564 71 - - 30 0.55 540 76 1.06 741 60 0.43 532 74 1.41 975 90 0.5 525 74 1.51 1082 120 0.4 520 75 1.61 1162 150 0.43 515 73 1.61 1181

Wet Cake + Paper + Filter: 172.4 g

Dry Cake + Paper + Filter: 108.46 g

Paper + filter paper: 11.5 g

Obtained dry cake: 96.96 g

ICP analysis results for arsenic, copper and iron in the recovered solids are summarized in Table 13 below.

As% Cu% Fe% As% Extraction Concentrate 3.22 0.09 8.38 0.0 Leaching 1 1.07 0.31 5.22 71.1 Leaching 2 0.25 0.30 2.85 92.7 Leaching 3 1.57 0.30 5.40 56.2

The above results clearly show that the reaction rate is significantly faster at 100 ° C than at 80 ° C.

Example 8

How to remove iron / arsenic

Way

The liquid (10 L) obtained from the pre-oxidation reaction was returned to the vessel and heated to 80 ° C. with gentle stirring. When the temperature was reached, the pH and Eh of the liquid were measured and samples were taken. The solution was then vented under stirring (100 L / hr), the pH and Eh of the liquid were measured and then samples were taken every 30 minutes. After 4 hours, the removal reaction was deemed complete, the liquid was filtered using a Buchner apparatus and the precipitate removed was removed as a filter cake. The wet cake was weighed and then dried in an oven for 24 hours. The dried cake was then weighed and the sample immersed for analysis.

Results and Discussion

The pH and Eh of the liquid over time are summarized in Table 14 below.

Minutes pH Eh Sample number As (g / L) Fe (g / L) AASCu (g / L) AASFe (g / L) Reference 0 0.7 500 One 2.8 3.7 88 3.7 Volume: 10L 0 0.7 500 - - - - - Air 100L / hour 30 1.1 510 2 2.3 3.6 90 3.3 60 1.5 520 3 1.2 2.7 90 2.6 90 1.6 525 4 0.6 2.4 91 2.2 120 2.0 530 5 0.3 1.3 85 1.6 150 2.1 535 - - - - - 180 2.1 545 - - - - - 210 2.2 555 - - - - - 240 2.5 570 6 ND 0.5 88 <0.1 Cu ^{total amount} 88, Cu ²⁺ 88

Wet Cake + Paper + Filter: 257.2 g

Dry Cake + Paper + Filter: 128.94 g

Paper + filter paper: 11.5 g

Obtained dry cake: 117.44 g

The analysis results of the precipitation are shown in Table 15 below.

element density (weight%) As 19.0% Fe 33.8% Cu 1.5%

Over four hours of experiment, almost 100% of iron and arsenic precipitated, while at the same time the oxidation potential (Eh) was returned to a high enough level for use in subsequent leaching procedures. The Fe / As molar ratio was 2.4, which can be expected to precipitate FeAsS with other Fe-containing compounds.

Example 9

How to remove iron / arsenic

Way

The liquid (10 L) obtained from the pre-oxidation reaction was returned to the vessel and heated to 80 ° C. with gentle stirring. When the temperature was reached, the pH and Eh of the liquid were measured and samples were taken. The solution was then vented under stirring (100 L / hr), the pH and Eh of the liquid were measured and then samples were taken every 30 minutes. After 4 hours, the removal reaction was deemed complete, the liquid was filtered using a Buchner apparatus and the precipitate removed was removed as a filter cake. The wet cake was weighed and then dried in an oven for 24 hours. The dried cake was then weighed and the sample immersed for analysis.

Results and Discussion

The pH and Eh of the liquid over time are summarized in Table 16 below.

Minutes pH Eh Sample number As (g / L) Fe (g / L) AASCu (g / L) AASFe (g / L) Reference 0 0.7 500 One 2.8 3.7 88 3.7 Volume: 10L 0 0.7 500 - - - - - Air 100L / hour 30 1.1 510 2 2.3 3.6 90 3.3 60 1.5 520 3 1.2 2.7 90 2.6 90 1.6 525 4 0.6 2.4 91 2.2 120 2.0 530 5 0.3 1.3 85 1.6 150 2.1 535 - - - - - 180 2.1 545 - - - - - 210 2.2 555 - - - - - 240 2.5 570 6 ND 0.5 88 <0.1 Cu ^{total amount} 88, Cu ²⁺ 88

Wet Cake + Paper + Filter: 257.2 g

Dry Cake + Paper + Filter: 128.94 g

Paper + filter paper: 11.5 g

Obtained dry cake: 117.44 g

Water content: 128.26 g (47.8%)

Example 10

Leaching by Treated Liquid with Low Regenerated Slurry Density

Way

A wet concentrate 90 g sample (about 24% water content, 68 g in terms of dry concentrate) was added to the treatment liquid (1.5 L) obtained from the oxidation reaction, and the formed slurry was stirred at 100-105 ° C. The pH and Eh of the liquid were monitored and samples were taken every 30 minutes over 4 hours. After this period has elapsed, the liquid is filtered using a Buchner apparatus and the filter cake is removed, and the wet cake is weighed and then dried in an oven for 24 hours. The dried cake was then weighed and the sample immersed for analysis.

Results and Discussion

The pH, Eh and component contents of the liquid over time are shown in Table 17 below.

Minutes pH Eh (mV) Sample number AASCu (g / L) AASFe (g / L) ICPAs (g / L) ICPCu (g / L) ICPFe (g / L) 0 2.5 600 - - - 10 1.8 555 - - - 30 1.5 550 One 90 1.5 60 1.3 545 2 89 1.1 90 1.2 540 3 87 1.2 120 1.1 540 4 67 0.8 150 1.1 535 5 88 1.0 180 1.0 535 6 85 1.0 210 1.0 535 7 88 1.0 240 1.0 536 8 87 1.1

Obtained dry cake: 58 g

First and Second Leaching Stage Embodiments

Example 11

First stage leaching

The purpose of this example is to determine if the operating conditions used for batch testing can be applied to commercial scale work through simulation of a continuous method. In addition, this experiment provides a material used for the oxidation of pyrite under atmospheric pressure.

As shown in FIG. 8, when continuously working under conditions similar to the batch process of Examples 5-7, As extraction efficiency gave results that are compatible with 85%.

step

A 7.5 liter titanium reactor was used, with a second reactor connected upstream from the first reactor, with a storage tank connected upstream therefrom. In a continuous operation, 2 liters / hour of solution was fed from the feed tank to the first reactor using a peristaltic pump. The solid addition rate was 144 g / h, which was obtained by batchwise adding 24 g (dry weight basis) of concentrate to the first reactor every 10 minutes.

Initially, a 30 liter raw material solution containing 80 g / l Cu ²⁺ , 200 g / l NaCl, 100 g / l CaCl ₂ and having a pH <1 was prepared. To each reactor, 7.5 liters of raw solution was added and maintained at 100 ° C., a P80 = 30 micron low grade gold (30 g / ton) concentrate with a dry equivalent of 360 g was added, followed by stirring the formed slurry. And every 30 minutes for Eh, pH, As, Fe and Cu. After 3 hours, 100 ml of slurry sample was taken and backwashed with Buchner funnel and washed with acid brine solution. The solid was then dried and analyzed for copper, arsenic and iron by ICP.

After 3 hours, continuous operation was performed for another 10 hours (according to the procedure as described above), and 200 ml of slurry sample was taken every 2 hours and filtered as described above. The solid was then dried and analyzed for copper, arsenic and iron by ICP. The results are shown in Tables 18-20.

number Solid sample Minutes Sample number T (℃) Eh (mV) pH Cu (g / l) Fe (g / l) As (g / l) Reference 01 0 0 95 620 One 77 0 0 720g dry solids added 2 60 One 100 520 0.5 87 2.6 2.3 All additions below are weights of wet concentrate 3 120 2 100 520 0.5 83 2.7 2.3 Add 192.95 g over 1 hour 4 180 3 90 530 0.7 81 1.7 1.4 Add 187.9 g over 1 hour 5 240 4 98 535 0.6 88 1.9 1.6 Add 183.4 g over 1 hour 6 300 5 103 540 0.4 89 2.5 2.1 Add 189.65 g over 1 hour 7 One 360 6 109 530 0.2 74 2.2 1.9 Add 195.44 g over 1 hour 8 420 7 109 535 0.2 95 3.7 3.3 Add 194.42 g over 1 hour 9 2 480 Instability Add 31.61 g over 1 hour 10 540 8 106 525 0.2 95 4.9 4.5 Add 200 g over 1 hour 11 600 9 102 524 0.3 95 4.2 3.8 Add 198.79 g over 1 hour 12 3 660 10 70 509 0.7 98 3.2 2.6 Add 62.82 g over 1 hour 13 4 680 11 74 501 0.8 99 3 2.4

number time Minutes Sample number T (℃) Eh (mV) pH Cu (g / l) Fe (g / l) As (g / l) 01 0 0 80 620 2 60 One 85 530 0.9 77 1.3 One 3 120 2 86 530 0.8 81 1.5 1.2 4 180 3 90 530 0.7 84 1.7 1.5 5 240 4 88 530 0.6 86 1.8 1.6 6 300 5 85 525 0.5 91 2.1 1.8 7 360 6 85 520 0.4 95 2.2 2 8 420 7 85 520 0.4 105 2.7 2.3 9 480 An unstable state 10 540 8 84 519 0.4 99 2.8 2.5 11 600 9 83 514 0.4 98 3.2 2.9 12 660 10 83 516 0.4 98 3.5 3.1 13 680 11 83 512 0.4 99 3.3 2.94

Solid analysis results for feedstock and tank 2 upstream number Explanation Fe As Cu As extraction Mass loss (%) (%) (%) (%) (%) Feedstock 8.60 3.20 0.08 0.00 0.00 One Tank 2-Solid 1 4.25 0.58 0.50 83.1 6.55 2      Solid 2 4.15 0.42 0.52 87.8 6.79 3     Solid 3 4.55 0.51 0.55 85.1 6.27 4     Solid 4 4.15 0.51 0.42 85.1 6.80 mixture     Solid 5 4.50 0.52 0.42 84.8 6.44 Average 85.2 6.57

Example 12

Second stage leaching

The purpose of this example is to assess the possibility of oxidizing the pyrite component of the residue obtained from the As leaching step using pure oxygen under atmospheric pressure. 500 g of the residue obtained during the continuous leaching test of Example 11 was used in this experiment.

Pyrite was continuously oxidized to oxygen at 105 ° C. under atmospheric pressure. The final As and Fe extraction rates were both 95% or more. S _(e) (elemental sulfur) in the oxidation residue was comparable to sulfur combined with arsenite in the concentrate. The results are shown graphically in FIG.

step

A 7.5 liter titanium reactor was prepared with a dispersive turbine stirrer and a suitable titanium gas injector on a large yellow hotplate. 5 L saline solution with the following composition was prepared in a 7.5 L titanium reactor: 250 g / l NaCl, 50 g / l CaCl ₂ , 20 g / L Cu (ferric chloride), adjusted to pH <1.0.

Representative samples of the residues of the As leaching step from Example 11 were fed to an external laboratory to analyze for elements S, total S, Au, Fe and As.

The stirrer run was set to 80 Hz on VSD, the temperature of the solution was raised to 105 ° C., samples were taken at t = 0, the Eh and pH were monitored and 500 g of the dry arsenic leaching residue obtained in Example 11 were introduced into the solution. Injected. After 30 minutes, a sample of the solution was taken and analyzed for Fe, As, Cu and Eh and pH were monitored.

Oxygen was introduced at a rate of 1 L / min. Eh, pH, Fe, Cu and As were monitored every 30 minutes for the first 3 hours and thereafter every 1 hour. When the soluble Fe assay stopped increasing, the experiment was considered complete.

The final sample was taken, the suspension filtered and the cake washed twice with acidic brine and then with hot water until the filtrate became clear. The washed cake was dried, weighed and analyzed for As, Fe, Cu, C, S in elemental state, and total S.

Elemental analysis

The following elemental analysis results were obtained.

Final leaching stage solid weight

cake:

Wet Cake + Paper + Filter: 744.27 g

Dry cakes + paper + filter paper: 490.96 g

Paper and filter paper: 6.09 g

Obtained dry cake: 424.87 g

Mass Reduction Rate: 15%

Solid Elemental Analysis of Feedstock for Oxidative Pyrite Oxidation

 Information on dry weight or ppm Explanation Fe (%) As (%) Cu (%) Au (ppm) S _t (%) S _e (%) Sum 4.31 0.55 0.61 - - - Availability 0.01 0.03 0.10 - - - Insoluble 4.30 0.51 0.51

Solid Elemental Analysis of Residues Obtained from Pyrite Oxidation by Oxygen

Information on dry weight or ppm Explanation Fe (%) As (%) Cu (%) Au (ppm) S _t (%) S _e (%) Sum 0.52 0.19 0.35 - - - Availability 0.01 0.01 0.01 - - - Insoluble 0.51 0.18 0.34 - - -

Extraction rate of Fe and As from concentrates compared to pyrite oxidation residue

As shown in Tables 23-26 below, the extraction rates of As and Fe both exceeded 95%.

Pyrite oxidation, feedstock Vs residues:

weight% Mass (g) Feedstock Residue Feedstock Residue 500 424.9 Fe 4.3 0.51 21.6 2.2 As 0.5 0.18 2.6 0.8

Corresponding concentrate Mass (g) density(%) 535.2 100 As 17.1 3.2 FeAsS 37.2 7.0 Fe 46.0 8.6 FeS ₂ 25.6 4.8 S 29.4 5.5 S in FeAsS 7.3 1.4 FeS ₂ S 22.1 4.1 Fe in AsFeS 12.8 2.4 Of FeS ₂ Fe 19.3 3.6 Other Fe 14.0 2.6

Extraction of Fe and As from Concentrates on Pyrite Oxidation Residues Fe 95.3% As 95.5%

Extraction of elemental sulfur from the pyrite oxidation residue showed that S (element) was equivalent to S combined with AsFeS, ie 1.4% in concentrate.

Elemental sulfur extraction Float head sample weight 8.68 g 100% S (Elemental) Extraction Residue Sum 0.35 g carbon 0.21 g S (element) 0.14 g 1.6% Mass change 79% S (element) related to concentrate 1.28%

Although the preferred method has been illustrated above, those skilled in the art will appreciate the other advantages of the present invention as described below.

The present invention provides the following advantages:

Using the method of the present invention, it is possible to recover noble metals from sulfur-containing ores and concentrates which are difficult or impossible to process by conventional methods / techniques such as smelting and roasting.

-The method of the present invention can be applied even when the carbon content in the ore is high. Since the method of the present invention is carried out in a solution, it is possible to prevent the adsorption of the precious metal on carbon which may interfere with the recovery of the precious metal using a blinding agent. Because there is.

Using the method of the present invention, contaminants can be removed from a wide range of ore and concentrate feedstocks and once contaminants are removed they can be treated using conventional smelting / roasting techniques.

-According to the method of the present invention, arsenic, iron, and sulfur can be removed from the original arsenite concentrate in an easy-to-dispose form, thereby providing a concentrate that is easy to treat.

The process of the present invention offers the possibility of recovering a wide range of economically valuable metals, especially precious metals, using simple bicyanide based leaching and separation processes including activated carbon adsorption.

When the method of the present invention is used for contaminated residues, the contaminated residues can later be disposed of so as to have less environmental impact.

Although the present invention has been described in detail based on the preferred embodiments, it should be understood that the protection scope of the present invention is not limited thereto, and various modifications and variations of the present invention can be made.

Claims (39)

As a method for recovering precious metals from sulfur-containing soils, Preparing an aqueous acid halide solution having an oxidation potential sufficient to oxidize sulfur-containing soils and solubilize the precious metal in solution; Adding the sulfur-containing soil to the aqueous acid halide solution to oxidize the sulfur-containing soil and solubilize the precious metal; And -Separating said precious metal from oxidized sulfur-containing soils How to include. A method for recovering precious metals from sulfur-containing soils contaminated with arsenic, Preparing an aqueous acid halide aqueous solution having a oxidation potential sufficient to oxidize the sulfur-containing soil and solubilize the precious metal in solution and to precipitate arsenic; Adding the sulfur-containing soil to the aqueous acidic halide solution to oxidize the sulfur-containing soil, solubilize the precious metal and precipitate the arsenic; And Separating said precious metal from oxidized sulfur-containing soil and precipitated arsenic; How to include. The method according to claim 1 or 2, The solution containing the noble metal is separated from the oxidized sulfur soil and precipitated arsenic (if present) in the solid-liquid separation step, and then the noble metal is recovered from the solution in the metal recovery step. The method of claim 3, wherein In the metal recovery step, adsorbing said noble metal on activated carbon in at least one carbon containing column. The method of claim 4, wherein After adsorbing the noble metal onto the activated carbon, carbon is eluted with a cyanide solution and the eluted product is fed to an electrolytic step for recovering the noble metal. The method according to any one of claims 3 to 5, Said metal recovery step is provided after said solid-liquid separation step in in-line form and prior to recycling said solution to a sulfur-containing soil oxidation step. The method according to any one of claims 1 to 6, The precious metal to be recovered is characterized in that the gold, silver, platinum or other platinum group metal. The method according to any one of claims 1 to 7, The aqueous halide solution is a soluble metal halide solution having a halide concentration of about 8 moles / liter. The method of claim 8, The halide is chloride or a halide mixture comprising chloride and bromide. The method according to claim 8 or 9, The metal in the dissolved metal halide solution is copper and / or iron and acts as a multi-valent species. The method according to any one of claims 1 to 10, The sulfur-containing soil oxidation step is the following step, (i) for uncontaminated single refractory pyrite soils, a sulfur-containing soil oxidation step comprising a single leaching step of oxidizing pyrite and solubilizing the precious metal; or (ii) in the case of contaminated single or double refractory pyrite soils, a sulfur-containing soil oxidation step comprising a two-step leaching process for supplying the solution obtained from the first leaching step to a second leaching step; A method comprising the one or more leaching steps, such as: The method of claim 11, In the case of (ii), the pyrite is arsenic (Arsenopyrite), and in the first leaching step, the oxidation potential is adjusted to leach arsenic into the solution, and once arsenic is leached, the arsenic is in the form of ferric arsenate. Adjusting the pH of the solution to precipitate, and in the second leaching step, adjusting the pH of the solution so that the arsenic is discharged from the process together with the oxidized sulfur-containing soil by leaching the pyrite component and maintaining the arsenic in the form of ferric arsenate precipitation. How to feature. The method of claim 12, In the first leaching step, at a pH of a solution of no more than 1 but not greater than about 0.5 to precipitate arsenic immediately after its leaching, under about 0.7-0.8 volts Eh sufficient to leach the contaminant and solubilize the precious metal. At a solution temperature of about 80-105 ° C., contacting the soil with the solution. The method according to claim 12 or 13, In the second leaching step, the soil is contacted with a solution having an Eh of about 0.8-0.9 volts sufficient to leach pyrite, the pH of the solution being less than or equal to 1 and greater than about 0.2 to precipitate arsenic immediately after its leaching. And the temperature of the solution is about 90 ° C to 105 ° C. The method according to any one of claims 1 to 14, Immediately after recovering the noble metal, the concentration of the chemical species in the process is controlled by precipitating iron sulfate using a solution condition adjusting step. The method of claim 15, In the solution conditioning step, limestone and calcium carbonate are added to the solution to form a hematite / gypsum precipitate, which is then filtered and treated with the solid residue discharged from the leaching step (s). Way. The method according to any one of claims 1 to 16, When a high concentration of carbon is present in the sulfur-containing soil, a surfactant is added during the sulfur-containing soil oxidation step to prevent the precious metal from adsorbing on the carbon in the soil, or activated carbon is added to the solution during the sulfur-containing soil oxidation step to remove the precious metal. Adsorbing onto activated carbon preferentially. The method of claim 17, Wherein said surfactant is at least one organic solvent, including kerosene or phenol ether. A method for removing contaminants from contaminated sulfur soils, The sulfur species are mixed in an aqueous solution to oxidize the contaminants with a polyvalent species having a relatively high oxidation state solubilizing the contaminants in solution to produce the sulfur-free soils free of contaminants. Reducing to an oxidized state; And Regenerating the polyvalent species to its relatively high oxidation state while simultaneously removing contaminants from the solution; Method comprising a. A method for removing contaminants from contaminated sulfur soils, Mixing sulfur-containing soils in an aqueous solution having an oxidation potential controlled to substantially oxidize only the pollutants to solubilize them in solution to produce contaminant-free soils; And Separating the solution from the contaminated desulfurized soil; Method comprising a. The method of claim 19 or 20, The contaminant is removed from the solution by precipitation in a separate precipitation step by introducing an oxidant into the solution. The method of claim 21, In the precipitation step, the pH of the solution is maintained at about pH 1.5-3. The method according to any one of claims 19 to 22, The contaminant is oxidized through a two-step leaching process to leach into the solution, wherein in the first leaching step, the oxidation potential is adjusted to oxidize substantially only the contaminant and solubilize it in the solution. Increasing the oxidation potential to oxidize sulfides in the removed soil. The method of claim 23, The contaminated soil is separated from the solution after the first leaching step and fed to a second leaching step, the solution being polluted after each leaching step to remove contaminants from the solution by precipitation. Separating from the removed soil. The method of claim 23 or 24, The sulfur-containing soil is pyrite, and in the first leaching step, the pollutant is oxidized at an temperature of about 105 ° C. or less under an Eh of a solution sufficient to oxidize the pollutant in an acidic aqueous solution having a pH of 1 or less, but not to substantially elute pyrite. And wherein the second leaching step oxidizes the pyrite soil in an acidic aqueous solution at a temperature of about 105 ° C. or less under Eh of a higher solution below pH 1 but sufficient to elute the pyrite. The method of claim 25, In the second leaching step, an oxidant such as oxygen, air, chlorine gas, hydrogen peroxide is added to the solution. The method according to any one of claims 19 to 26, The solution is recycled throughout the process, wherein the solution is a dissolved metal chloride solution having a chloride concentration of about 8 mol / liter, wherein the metal in the dissolved metal chloride solution acts as a polyvalent species. The method according to any one of claims 19 to 27, The sulfur-containing soil has a high concentration of carbon, and during the pollutant oxidation step, a surfactant is added to the solution to prevent the precious metals from adsorbing on the carbon in the soil. The method according to any one of claims 19 to 28, And at least one metal recovery step for recovering metals leached into solution with contaminants and / or metals present in the soil from which residual contaminants have been removed. The method of claim 29, The sulfur-containing soil is a double refractory ore containing carbon, and after the final leaching step, a metal recovery step is performed to recover metal present in the soil which has been removed from the contaminants adsorbed on the carbon. . The method of claim 30, The metal recovery step includes a conventional roasting or smelting step, optionally leaching by chlorine or cyanide to recover the metal remaining in the roasted solid soil after the roasting or smelting step. Method for carrying out the. The method of any one of claims 29-31, Inline metal recovery step prior to and / or after the contaminant precipitation step to remove the leached metal into solution in the leaching step. The method according to any one of claims 29 to 32, Before recovering the metal, a plurality of soil separation steps are carried out to separate the soil from which the contaminants have been removed from the solution. The method of claim 33, wherein Wherein the separated solution after the leaching step or after each leaching step is fed to the contaminant recovery step, while the separated refined soil is fed to the metal recovery or disposal step. The method according to any one of claims 23 to 34, wherein After the contaminant precipitation step, after carrying out the contaminant separation step to remove the contaminants from the solution, the solution is recycled to the leaching step. A method for recovering precious metals in soil by treating contaminated sulfur soils having a relatively high carbon content, Leaching the sulfur-containing soil in an aqueous solution to leach the metal into the solution, during which the carbon in the soil is shielded to prevent the precious metal from adsorbing on the carbon; And Recovering the noble metal from the solution. The method of claim 36, Wherein said carbon is masked using a surfactant as defined in claim 18. 38. The method of claim 36 or 37, 36. A method as defined in any one of claims 1 to 35. A metal produced by the method as defined in any one of claims 1 to 38.